GIFT   OF 


HYDROMETALLURGY 
OF    SILVER 


WITH    SPECIAL    REFERENCE    TO 

CHLORIDIZING    ROASTING    OF   SILVER   ORES   AND  THE 

EXTRACTION  OF  SILVER  BY  HYPOSULPHITE 

AND   CYANIDE   SOLUTIONS 


BY 
OTTOKAR    HOFMANN 

Mining  and  Metallurgical  Engineer,  Member  of  the  American  Institute 

of  Mining  Engineers,  of  the  American  Chemical  Society 

and  of  the  American  Electrochemical  Society 


NEW  YORK  AND  LONDON 

HILL    PUBLISHING    COMPANY 

1907 


T~/V 


Copyright,  1907,  BY  HILL  PUBLISHING  COMPANY. 

ALSO  ENTERED  AT  STATIONERS*  HALL,  LONDON,  ENGLAND 


All  rights  reserved 


.     , .     »,    «,„,,, 


:  -:.:•  '• 


Hill  Publishing  Company  New  York  U.S.A. 


PREFACE 

THE  silver  ores  which  are  treated  by  a  hydrometallurgical 
process  are  mostly  complex  sulphide  ores  consisting  of  quite  a 
number  of  different  metal  sulphides.  In  order  to  render  soluble 
the  silver  contained  therein  the  ore  is  roasted  with  an  addition 
of  salt  (chloridizing  roasting),  by  which  process  the  silver  is 
converted  into  silver  chloride.  In  this  chemical  reaction,  how- 
ever, all  or  nearly  all  the  constituent  minerals  of  the  ore  partici- 
pate, which  makes  the  process  rather  complicated  and  we  may 
even  say  delicate,  because  the  formed  metal  chlorides  are  volatile 
and  induce  the  silver  chloride  to  volatilize  too,  and  in  order  to 
keep  this  loss  at  the  minimum  great  care  and  skill  are  required. 
The  solvent,  whether  this  be  sodium  hyposulphite  or  concen- 
trated brine,  will  extract  all  the  silver  which  was  converted  into 
chloride,  and  the  final  result  of  extraction  depends,  therefore, 
entirely  on  the  quality  of  the  roasting.  This  being  the  case,  it  is 
obvious  that  chloridizing  roasting  is  the  most  important  part  of 
the  process,  and  that  a  metallurgist  can  not  expect  to  be  success- 
ful in  the  hydrometallurgy  of  silver  unless  he  has  a  thorough 
knowledge  of  chloridizing  roasting  and  the  ability  to  apply 
skilfully  this  knowledge  in  actual  practice.  For  this  reason  a 
large  part  of  this  treatise  is  devoted  to  the  art  of  chloridizing 
roasting,  which  I  consider  to  be  especially  necessary,  as  there  is 
no  new  literature  on  this  subject,  though  great  advance  has  been 
made  in  it. 

Chloridizing  roasting  was  originally  studied  and  practised  in 
relation  to  amalgamation.  In  amalgamation  not  only  the  silver 
amalgamates,  but  the  base-metal  chlorides  also  amalgamate, 
which  greatly  interferes  with  the  process,  causing  a  poor  extrac- 
tion, a  great  loss  of  silver  and  mercury,  and  the  production  of 
very  base  bullion.  To  avoid  these  difficulties  and  to  make  the 
process  applicable  to  a  greater  variety  of  ores,  these  objectionable 
metal  chlorides  are  partly  expelled  and  partly  converted  into 

iii 

337632 


iv  PREFACE 

oxides  by  increasing  the  temperature  of  roasting.  The  expulsion 
of  the  volatile  metal  chlorides,  however,  induces  quite  a  percent- 
age of  the  otherwise  not  volatile  silver  chloride  to  volatilise, 
thus  causing  a  considerable  loss  of  silver. 

I  was  the  first  to  introduce  the  process  of  lixiviation  with 
sodium  hyposulphite  in  Mexico,  in  1868,  and  made  a  special 
study  in  actual  practice  of  chloridizing  roasting,  and  in  course 
of  time,  and  as  experience  was  gained  with  different  ores,  became 
convinced  that  chloridizing  roasting  as  practised  for  amalga- 
mation was  not  the  proper  way  to  roast  for  lixiviation.  A  large 
percentage  of  base-metal  chlorides  in  the  roasted  ore  is  detrimental 
to  a  successful  extraction  of  the  silver  by  amalgamation,  while 
in  lixiviation  they  do  not  interfere  with  the  extraction;  why  then 
should  we  expel  and  destroy  these  chlorides  by  increased  heat 
at  a  sacrifice  of  silver?  The  expulsion  of  the  volatile  compounds 
by  increased  heat  is  the  sole  cause  of  the  silver  loss  by  volatiliza- 
tion, and  if  we  avoid  this  we  consequently  will  avoid  this  losSj  or 
at  least  reduce  it  to  the  minimum.  I  therefore  modified  the  mode 
of  chloridizing  roasting,  inasmuch  as,  instead  of  expelling  the 
volatile  compounds  by  increased  heat,  I  endeavored  to  retain 
them  as  much  as  possible  in  the  roasted  ore  by  using  the  lowest 
permissible  temperature  —  a  temperature  sufficiently  high  to  pro- 
duce the  chemical  reaction  but  not  high  enough  to  expel  the  metal 
chlorides.  Of  course  such  a  condition  cannot  be  maintained  with 
theoretical  exactness,  but  I  succeeded  in  greatly  reducing  the 
loss  of  silver  by  volatilization  even  with  ores  rich  in  arsenic. 
By  this  modification  in  chloridizing  roasting  a  marked  step  forward 
was  made  in  the  hydrometallurgy  of  silver. 

The  second  part  of  this  treatise  is  devoted  to  the  extraction 
of  the  silver  from  the  roasted  ore  by  different  solvents.  The  last 
chapter  deals  with  the  cyaniding  of  silver  ores.  This  process,  so 
extensively  and  successfully  used  for  the  extraction  of  gold,  is 
still  more  or  less  in  its  experimental  stage  with  regard  to  silver 
ores.  However,  such  very  promising  results  have  been  obtained 
with  certain  ores,  that  further  experiments  and  a  thorough 
investigation  of  this  subject  are  to  be  recommended.  For  complex 
auriferous  silver  ores  a  combination  of  the  sodium  hyposulphite 
and  the  cyanide  processes  is  most  promising. 

By  far  the  larger  part  of  this  treatise  is  a  record  of  my  long 
years  of  experience,  studies,  and  experiments  on  a  commercial 


PREFACE  V 

scale  rather  than  the  product  of  compilation,  and  it  will  be  found 
that  attention  is  paid  to  many  things  apparently  unimportant 
but  which,  in  actual  practice,  I  have  found  to  be  of  great  impor- 
tance, and  on  the  correct  execution  of  which  the  success  of  the 
enterprise  may  often  depend. 

I  hope  that  this  treatise  will  prove  a  friend  and  adviser  to  the 
student  of  hydrometallurgy  as  well  as  to  the  metallurgist  in  active 
service  in  the  field. 

OTTOKAR  HOFMANN. 
KANSAS  CITY,  Mo.,  January,  1907. 


TABLE  OF  CONTENTS 


PREFACE     . 

TABLE  OF  CONTENTS 

LIST  OF  ILLUSTRATIONS 


PAGES 
iii 
vii 
ix 


PART  I.    CHLORIDIZING  ROASTING  OF  SILVER  ORES 

CHAPTER  I.    THEORY  OF  CHLORIDIZING  ROASTING 1-10 

Behavior  of  different  minerals  in  chloridizing  roasting,  5. 
Classification  of  ores  in  relation  to  chloridizing  roasting,  9. 

CHAPTER  II.     CRUSHING  OF  THE  ORE 'r       11-14 

CHAPTER  III.     PERCENTAGE  OF  SALT  REQUIRED        .     .     .;   .     .         15-19 

The  proper  time  to  add  the  salt,  16. 

CHAPTER  IV.     Loss  OF  SILVER  BY  VOLATILIZATION         ....         20-25 
Method  of  ascertaining  the  loss  of  silver  by  volatilization,  22. 

CHAPTER  V.     METHODS  OF  ROASTING 26-41 

Chloridizing  self-roasting,  26.  Chloridizing  heap-roasting, 
27.  Chloridizing  roasting  with  steam,  31.  Chloridizing  roast- 
ing of  silver  ores  containing  gold,  35.  Chloridizing  roasting 
for  amalgamation,  37. 

CHAPTER  VI.     CONSUMPTION  OF  FUEL .      .         42-44 

CHAPTER  VII.     REVERBERATORY  FURNACES  WORKED  BY  HAND     .         45-61 

The  single-hearth  reverberatory,  45.     The  two-story  single- 
hearth  furnace,  46.     The  long  reverberatory  furnace,  47.     The 
two-story  long  furnace,  60. 
CHAPTER  VIII.     MECHANICAL  ROASTING  FURNACES        ....         62-86 

Mechanical  furnaces  fed  by  charges,  63.  Mechanical  roasting 
furnaces  with  continuous  feeding,  71. 

CHAPTER  IX.    COLLECTING  THE  FLUE-DUST 87-93 

CHAPTER  X.    SULPHATING  ROASTING 94-98 

CHAPTER  XI.    CHLORIDIZING  OF  ARGENTIFEROUS  ZINC-LEAD  ORE      99-126 

Roasting  experiments,  101.  Roasting  in  the  Stetefeldt  fur- 
nace, 104.  Re  roasting  the  ore  from  the  shaft,  106.  Re  roasting 
the  ore  of  the  Stetefeldt  furnace  in  the  modified  Howell  furnace, 
109.  Application  of  steam,  110.  Conclusions,  111.  Roasting 
in  the  White-Howell  furnace,  112.  Roasting  in  the  modified 
Howell  furnace,  113.  Additional  chlorination  after  the  ore 
has  left  the  furnace,  115.  Results,  115.  Loss  of  silver  by 
volatilization,  117.  The  roasted  ore,  118.  Consumption  of 
wood,  119.  Cost  of  roasting  in  the  modified  Howell  furnace, 


Vlll 


TABLE  OF  CONTENTS 


PAGES 


119.  Summary,  120.  Roasting  in  the  reverberatory  furnace, 
121.  Oxidizing  roasting,  122.  Treating  the  oxidized  ore  with 
cupric  chloride,  125.  Consumption  of  wood  in  the  reverbera- 
tory furnace,  125.  Cost  of  roasting  in  the  reverberatory 
furnace,  126. 

CHAPTER  XII.    CHLORIDIZING  OF  CALCAREOUS  ORES       .... 

Roasting  in  the  Bruckner  furnaces,  128.     Adding  the  salt  in 

the  furnace,  135.     Adding  the  salt  in  the  battery;  self  roasting, 

140.     Balling  of  the  ore,  142.     Roasting  in  the  reverberatory 

furnaces,  144.     Conclusion,  151. 


127-152 


PART  II.    EXTRACTION  OF  THE  SILVER 

CHAPTER  XIII.     LIXIVIATION  WITH  SODIUM  HYPOSULPHITE     .         155-184 

Base-metal  leaching,  157.     Silver  leaching,  174. 
CHAPTER  XIV.     PRECIPITATION  OF  SILVER     .     .     ...     .     .     185-193 

CHAPTER  XV.    TREATMENT  OF  THE  PRECIPITATE 194-214 

CHAPTER  XVI.    CONSTRUCTION  OF  TROUGHS        .     .     .    '.     .     .     215-218 

CHAPTER  XVII.    TROUGH  LIXIVIATION 219-250 

The   troughs,    224.     Sluice-tanks   and   sluicing,    225.     Ar- 
rangement   and    operations,    229.     Precipitating    vats,    233. 
Practice  of  trough  lixiviation  at  Cusihuiriachic,  240.     Trough 
lixiviation  experiments  on  a  large  scale,  243.     Time  required 
for  base-metal  leaching,  244.     Quantity  of  water  required,  245. 
Quantity  of  silver  dissolved  by  the  base-metal  solution,  246. 
Silver   leaching,    246.     Quantity   of   solution    required,   248. 
Fineness    of    the    precipitate,    248.     Advantages    of    trough 
lixiviation,  249. 
CHAPTER  XVIII.    THE  RUSSELL  AND  Kiss  PROCESSES         .     .      .     251-255 

The  Russell  process,  251.     The  Kiss  process,  254. 
CHAPTER  XIX.    THE  AUGUSTIN  PROCESS        ...     ...     .     256-257 

CHAPTER  XX.     EXTRACTION  WITH  SULPHURIC  ACID        .  -  .     ..    .     258-280 

Extraction  of  silver  from  copper  matte,  258.     Extraction  of 
silver  from  black  copper,  278. 
CHAPTER  XXI.    THE  ZIERVOGEL  PROCESS      ....]..     .     281-282 

CHAPTER  XXII.    TREATMENT  OF  SILVER  ORES  RICH  IN  GOLD     .     283-286 
CHAPTER  XXIII.     CYANIDATION  OF  AURIFEROUS  SILVER  ORES       .     287-328 

Treatment  of  raw  ore,  287.     Cyaniding  auriferous  silver  ores 
at  Palmarejo,  Mexico,  288.     Cyaniding  auriferous  silver  ores  at 
San  Salvador,  C.  A.,  321.    Testing  the  cyanide  solution  for 
gold  and  silver,  326. 
INDEX  329-345 


LIST  OF  ILLUSTRATIONS 

FIGURES  PAGES 

1-2.  Single-hearth  reverberatory  furnace 46 

3.  Two-story,  single-hearth  reverberatory  furnace     .      .  47 

4-6.  Long  reverberatory  furnace         .      .      .      .      .      .      .  49-50 

7-8.  Plan  and  elevation  of  working  door 57 

9  A.  The  Kiistel  working  door 58 

9  B-9  C.  Device  for  working  door        ........  59 

10-12.  Long  reverberatory  furnace,  two-story       ....  60-61 

13-14.  Bruckner  roaster .  63 

15-17.  Hofmann  improved  Bruckner  furnace         ....  68 

18.  O'Harra  furnace 73 

19-20.  Horizontal  and  cross-section  of  Ropp  furnace        .      .  75 

21  A-21  B.     Ropp  furnace,  longitudinal  elevation  and  plan      .      .  76 

22.  Howell-White  furnace      .      ......     .      .      .  78 

23.  Howell  furnace,  discharge  end  and  ore-vault         .      .  79 

24.  Stetefeldt  furnace 84 

25.  Feeding  machine,  Stetefeldt  furnace 85 

26.  Vertical  section  of  Hofmann  dust  collector       ...  88 
27-28.  Details  of  bars  and  bearings,  Hofmann  dust  collector  89 

29.  Position  of  bars,  Hofmann  dust  collector.      ...  90 

30.  Horizontal  section,  Hofmann  dust  collector     ...  92 

31.  Leaching  tank,  vertical  section         .      .  "~7     .      .      .  158 

32.  Leaching  tank,  plan    ..........  159 

33-34.  Brass  clamps  for  \\-  and  2-in.  hose 161 

35.  Calcium  sulphide  plant     .      .      .      ...     .      .      .  187 

36.  Distributing  trough  for  milk  of  lime 188 

37-39.  Boiler  and  pressure  tank  for  calcium  sulphide       .      .  189-190 

40-41.  Air  blow-off  drum 200-201 

42.  Horizontal  pressure  tank,  for  solution         ....  201 

43.  Cast-iron  flange  union  for  discharge  pipe  of  pressure 

tank         .      .      . .      ..-         202 

44.  Apparatus  for  the  manufacture  of  lye        . .      .      .      .  204 

45.  Pressure  tanks  for  treatment  of  precipitate      .      .      .  206 
46-47.  Drying  and  roasting  furnace  for  silver  precipitate       .  208-209 
48.  Dust-collecting  afrangement  for  cupeling  furnace       ."  212 
49-51.  Trough:  cross  section,  connection,  union    ....  217 
52-54.  Settling-tank  arranged  for  sluicing         226-227 

55.  Wheel  for  closing  discharge  gate       ......  228 

56.  System  for  continuous  trough  lixiviation    ....  230 
57-58.  Precipitation  tank,  vertical  section  and  plan         .      .  234-235 

ix 


x  LIST  OF  ILLUSTRATIONS 

FIGURES  PAGES 

59.  Precipitating  vat         236 

60-61.  Filter  frame .      .  237 

62-63.  Lump-grinding  machine,  elevation  and  plan    .      .      .  238-239 

64-65.  Lump-grinding  machine,  mantle  and  muller    .      .      .  240 

66.  Stir  tank,  vertical  section      .      . 264 

67-69.  Cast-iron  pressure  tank    .      .      j      .   '.     ,,     .      .      .  266 

70.  Tower  for  refining  cupric  sulphate  solutions     .      .      .  268 

71-72.  Pan  evaporator,  longitudinal  vertical  section  and  plan  271 

73.  Device  for  discharging  blue  vitriol 276 

74-75.  Cyanide  leaching  plant,  plan  and  elevation      .      .      .  293-294 

76-77.  Plan  and  section  of  slime  plant 306-307 

78.  Agitation  vat   .      rf     .      .     ...      .      .     .      .      .      .  308 

79.  Decantation  vat     ./..<...         ....  311 

80.  Timber  foundations  supporting  decantation   vats   of 

slime  plant 312 

81.  Decantation  vats  in  course  of  construction       .      .      .  312 

82.  General  arrangement  of  slime  plant 314 

83.  Three  of  the  agitation  vats  and  top  of  two  of  the 

decantation  vats      .  314 


PART  I 
CHLORIDIZING    ROASTING    OF    SILVER   ORES 


THEORY  OF  CHLORIDIZING  ROASTING 

THE  object  of  chloridizing  roasting  is  to  convert  the  silver  in 
the  ore  into  silver  chloride,  in  which  state,  while  not  soluble  in 
water,  it  becomes  soluble  in  sodium  hyposulphite  and  other  solu- 
tions like  hot  concentrated  brine,  potassium  cyanide,  etc.,  by 
means  of  which  it  can  be  extracted  from  the  ores.  It  is  one  of 
the  most  complicated,  and  in  the  hydrometallurgy  of  silver  the 
most  important,  of  metallurgical  operations.  The  results  of  the 
subsequent  extraction  of  the  silver  by  the  solvent  depend  entirely 
on  the  quality  of  the  roasting.  Silver  chloride  dissolves  easily, 
and  even  a  very  dilute  solution  of  sodium  hyposulphite  will 
extract  all  the  silver  which  was  converted  into  chloride  during 
roasting,  so  that  it  is  of  the  greatest  importance  that  this  part  of 
the  process  be  executed  intelligently,  and  with  great  care  and  skill. 

The  ores  which  are  subjected  to  chloridizing  roasting  are 
usually  complex  sulphide  ores,  though  in  some  instances  ores 
almost  free  of  sulphides  are  roasted  successfully,  but  these  are 
exceptional  cases.  To  effect  chloridizing  roasting  chlorine  has 
to  be  generated  in  the  ore  while  being  subjected  to  heat.  This  is 
done  by  an  addition  of  salt  (sodium  chloride)  to  the  ore.  But 
not  only  the  silver  is  converted  into  a  chloride;  all  the  constit- 
uent parts  of  the  ore  also>  undergo  a  change,  quite  frequently  even 
the  gangue.  During  the  first  part  of  the  roasting  the  sodium 
chloride  remains  indifferent,  while  the  metal  sulphides  oxidize, 
forming  metal  sulphates  and  sulphurous  acid;  then  by  the  action 
of  these  sulphates  on  the  salt  rather  complicated  reactions  take 
place,  by  which  metal  chlorides,  chlorine,  hydrochloric  acid,  and 
sulphurous  chloride  are  formed. 

The  decomposition  of  the  sodium  chloride  and  the  chlorination 
of  the  silver  and  other  metals  is  effected  in  the  furnace  in  different 
ways: 

(1)  In  oxidizing  the  metal  sulphides  there  is  always,  besides 

3 


4  HYDROMETALLURGY  OF  SILVER 

sulphurous  acid  gas,  some  sulphuric  acid  gas  formed.  Not  so 
much  in  the  beginning  as  later,  when  part  of  the  metal  sulphides, 
especially  the  iron,  have  changed  into  oxides,  which  then  act  as 
a  contact  substance  on  the  sulphurous  acid,  converting  it 
into  anhydrous  sulphuric  acid,  which  then  decomposes  the  sodium 
chloride.  The  formation  of  sulphuric  acid  increases  much  if  a 
liberal  amount  of  air  is  permitted  to  enter  the  furnace,  and  as 
the  sulphuric  acid  plays  an  important  part  in  chloridizing  roasting, 
provision  should  be  made,  in  the  construction  of  the  furnaces, 
that  they  may  receive  as  much  air  as  required. 

(2)  By  the  reaction  between  metal  sulphates  and  the  sodium 
chloride,  by  which  metal  chlorides  and  sodium  sulphate  are  formed. 
This  is  the  principal  reaction  for  the  formation  of  chlorides.     The 
metal  sulphates  which  act  most  energetically  in  this  respect  are 
those  of  iron  and  copper,  for  which  reason  ores  containing  an 
ample  amount  of  iron  pyrites  and  some  copper  sulphides  will  be 
found  to  chloridize  the  best. 

(3)  Besides  chlorine,  there  is  also  hydrochloric  acid  formed, 
owing  to  the  moisture  in  the  air  and  fuel.     Hydrochloric  acid 
acts  very  energetically,   and  sometimes  it  is  of  advantage  to 
produce  larger  quantities  of  it,  in  which  case  steam  is  admitted 
into  the  furnace  to  supply  an  extra  amount  of  moisture. 

(4)  The  fumes  of  volatilized  salt  (sodium  chloride)  act  also  in 
chloridizing    the    ore.     Quartz    decomposes    the    salt,    forming 
s'licate  of  soda  and  chlorine,  but  it  takes  a  rather  high  heat  for 
this  reaction,  and  only  in  exceptional  cases  does  it  come  into  play. 

(5)  Volatile  metal  chlorides  act  also,  chloridizing  the  silver, 
whereby  they  are  reduced  to  subchlorides  or  changed  into  oxides. 
Cupric  chloride  acts  very  energetically  in  this  respect. 

If  salt  and  ore  are  charged  together,  the  salt  is  not  decomposed 
until  the  formation  of  sulphates  begins,  and  the  first  stage  in 
roasting  is,  therefore,  a  mere  oxidizing  process.  Whatever  sul- 
phuric acid  is  formed  during  this  period  by  the  oxidation  of  the 
sulphides  acts  more  readily  on  the  base  metals,  forming  sulphates, 
than  on  the  salt.  Likewise  it  acts  more  readily  on  the  lime  and 
other  earthy  matters  of  the  gangue.  The  principal  part  of  the 
chlorination  takes  place  by  the  reaction  between  the  metal 
sulphates  and  the  salt.  The  oxidizing  and  chloridizing  periods 
are  quite  distinct,  and  can  be  easily  observed  by  the  appearance 
of  the  ore  in  the  furnace  and  by  the  smell  of  the  fumes  of  a  sample 


THEORY  OF  CHLORIDIZING  ROASTING  5 

taken  from  the  charge.  During  the  oxidizing  period  the  glow  of 
the  surface  of  the  ore  is  much  brighter  than  the  inside,  and  the 
particles  brought  to  the  surface  by  stirring  brighten  instantly  to 
a  lighter  red.  The  fumes  of  a  sample  have  a  strong,  choking 
smell  of  sulphurous  acid.  During  the  chloridizing  period,  if  an 
excessive  fire  is  not  kept  up,  the  surface  of  the  ore  assumes  a  very 
dull  red,  while  the  deeper  layers  are  of  a  brighter  glow,  which, 
however,  becomes  dull  shortly  after  the  particles  are  brought 
to  the  surface.  The  fumes  of  a  sample  have  a  mild  but  distinct 
odor  of  chlorine.  The  charge  swells  and  becomes  loose  and 
woolly. 

It  was  mentioned  above  that  the  ores,  which  are  subjected  to 
chloridizing  roasting,  are  mostly  complex  argentiferous  ores,  and 
as  the  roasting  is  much  influenced  by  the  behavior  of  the  con- 
stituent parts  of  the  ore  and  has  to  be  modified  according  to 
the  requirements  of  one  or  the  other  of  the  constituents,  a 
knowledge  of  the  behavior  of  the  different  minerals  and  the 
changes  they  undergo  during  chloridizing  roasting  is  therefore  indis- 
pensable in  order  to  conduct  the  process  intelligently. 

BEHAVIOR  OF  DIFFERENT  MINERALS  IN 
CHLORIDIZING  ROASTING 

Iron  Pyrites.  —  During  the  oxidizing  period,  sulphurous  and 
sulphuric  acids  are  formed,  of  which  the  former  escapes  entirely, 
while  part  of  the  latter  combines  with  lime  and  other  earthy 
matters  of  the  gangue,  and  part  combines  with  the  iron,  forming 
sulphates.  The  iron  changes  into  ferrous  and  ferric  sulphates  and 
into  ferric  oxide.  The  ferrous  and  ferric  sulphates  act  on  the 
salt,  forming  ferrous  and  ferric  chlorides  and  sodium  sulphate, 
while  some  of  the  chlorine  combines  with  sulphur  to  form  sulphurous 
chloride,  which  escapes  as  gas.  In  the  course  of  the  process 
both  these  iron  chlorides  give  off  their  chlorine,  chloridizing  the 
silver  and  changing  into  ferric  oxide.  In  practice  the  reaction 
is  not  quite  so  complete,  and  in  the  finished  charge  we  find, 
besides  the  ferric  oxide,  some  ferric  sulphate  and  some  ferrous 
and  ferric  chloride. 

The  iron  chlorides  decompose  easily  and  act  as  the  principal 
chloridizers,  for  which  reason  it  is  very  desirable,  in  fact  often 
necessary,  to  have  iron  sulphides  in  the  ore. 


6  HYDROMETALLURGY  OF  SILVER 

Copper  pyrites  consists  of  the  sulphides  of  copper  and  iron. 
During  oxidizing,  cupric  sulphate,  cuprous  and  cupric  oxides 
are  formed.  The  cuprous  oxide,  however,  soon  changes  into 
cupric.  During  chloridizing,  sulphurous  chloride  (which  vola- 
tilizes), cupric  and  cuprous  chlorides  are  formed.  Both  these 
copper  salts  melt  below  red  heat,  and  are  absorbed  by  the  ore, 
thus  becoming  finely  divided  through  the  ore  and  coming  in  inti- 
mate contact  with  the  silver.  Both  are  volatile.  At  a  higher 
heat  the  cupric  chloride  gives  off  part  of  its  chlorine,  chloridiz- 
ing the  silver  and  changing  into  cuprous  chloride.  In  presence 
of  steam,  hydrochloric  acid,  cuprous  and  cupric  oxides  are 
formed. 

The  iron  sulphide  of  the  copper  pyrites  undergoes  the  same 
chemical  changes  as  the  iron  pyrites.  For  this  reason,  and  for 
the  fact  that  cupric  chloride  gives  off  part  of  its  chlorine,  copper 
pyrites  is  a  very  good  producer  of  chlorine  during  roasting. 

In  the  roasted  charge  we  find  cupric  oxide,  cupric  sulphate, 
ferric  oxide,  ferric  sulphate,  ferrous  and  ferric  chlorides,  and 
cuprous  and  cupric  chlorides.  If  such  a  charge  is  subjected  to  a 
prolonged  roasting  at  a  high  heat  (dead  roast),  all  the  iron  as 
well  as  the  copper  will  be  changed  into  oxide. 

Other  copper  ores,  like  gray  copper  ore,  fahlerz,  etc.,  undergo 
the  same  changes. 

Galena  (lead  sulphide)  undergoes  the  changes  much  slower. 
It  cakes  easily,  for  which  reason  the  temperature  in  the  begin- 
ning has  to  be  kept  low  until  most  of  its  sulphur  has  been  oxidized. 
During  oxidizing,  sulphurous  acid,  lead  oxide,  and  lead  sulphate 
are  formed.  The  lead  sulphate  does  not  decompose  the  salt 
at  a  roasting  heat  and,  therefore,  does  not  take  an  active  part  in 
the  generation  of  chlorine.  When  air  has  free  access,  most  of 
the  lead  is  converted  into  sulphate  and  but  little  into  chloride, 
while,  if  the  supply  of  air  is  limited,  much  more  lead  chloride 
is  formed,  and  thus  becomes  a  consumer  of  chlorine.  It  is  volatile, 
and  volatilizes  without  giving  off  any  chlorine.  Lead  oxide  is 
volatile  too,  while  the  sulphate  remains  more  indifferent.  In 
the  roasted  ore  we  find  lead  sulphate  and  lead  chloride,  but  much 
less  of  the  latter. 

By  the  above  it  can  be  seen  that  lead  sulphide  is  not  a  desir- 
able constituent  part  of  a  roasting  charge. 

Zinc  Blende.  —  During  the  oxidizing   periods   zinc   sulphide 


THEORY   OF  CHLORIDIZING   ROASTING  7 

changes  into  zinc  oxide  and  zinc  sulphate,  but  the  sulphate  does 
not  act  decomposingly  on  the  salt.  By  the  action  of  the  chlorine 
and  hydrochloric  acid  zinc  chloride  is  formed,  which  is  very 
volatile  and  goes  off  in  heavy  fumes,  which  increase  when  the 
temperature  is  raised.  These  escaping  fumes  induce  the  silver 
to  volatilize,  for  which  reason  ores  rich  in  zinc  blende  have  to  be 
roasted  at  a  low  heat  to  avoid  an  excessive  loss  of  silver. 

In  the  roasted  ore  we  find  principally  zinc  oxide,  then  zinc 
sulphate  and  chloride. 

Zinc  blende,  as  a  rule,  contains  more  or  less  iron  sulphide, 
some  of  its  varieties  as  much  as  22  and  even  28  per  cent.  The 
iron  sulphide  takes,  of  course,  an  active  part  in  the  generation 
of  chlorine;  still  it  takes  much  skill  to  chloridize  satisfactorily 
the  silver  contained  in  zinc  blende.  This  subject  will  be  treated 
exhaustively  in  another  chapter. 

Arsenical  Pyrites.  —  This  consists  of  arsenic  sulphide  and 
iron  sulphide.  Arsenic  is  very  volatile  and  begins  to  come  off  from 
the  ore  in  dense  fumes  right  at  the  beginning  and  before  other 
sulphides  are  ignited.  During  this  part  of  the  process  much 
arsenate  of  silver  is  formed,  up  to  50  and  54  per  cent,  of  the  total 
silver  contained  in  the  ore.  This  silver  compound  is  soluble  in  a 
solution  of  sodium  hyposulphite.  During  the  chloridizing  period, 
however,  most  of  it  is  decomposed  without  volatilizing,  if  the 
temperature  is  kept  low,  but  it  volatilizes  very  readily  at  a  high 
heat,  causing  a  heavy  loss  in  silver.  Such  ores  have  to  be 
roasted  at  a  very  low  heat.  This  subject  is  exhaustively  treated 
in  another  chapter. 

In  roasting  arsenical  pyrites,  arsenious  oxide,  sulphurous  chlo- 
ride, arsenic  chloride,  and  ferric  chloride  are  formed  and  volatilized. 
In  the  roasted  charge  we  find  ferric  oxide,  ferric  sulphate,  ferrous 
and  ferric  chlorides,  and  some  ferric  arsenate. 

Antimony  Sulphide.  —  This  mineral  we  find  quite  frequently 
in  complex  silver  ores,  and  if  it  occurs  in  large  quantities  the 
roasting  has  to  be  conducted  very  carefully  and  at  a  very  low 
heat  on  account  of  its  great  volatility,  which  can  cause  a  heavy 
loss  of  silver.  During  oxidizing  it  changes  to  oxide  of  antimony, 
of  which  a  large  portion  is  volatilized  as  such.  During  chloridiz- 
ing antimony  trichloride  and  sulphurous  chloride  are  volatilized. 
In  the  roasted  ore  we  find  the  antimony  as  antimony  anti- 
monate. 


8  HYDROMETALLURGY  OF  SILVER 

Quartz.  —  We  find  quartz  quite  frequently  as  gangue  of  the  ore. 
At  a  proper  roasting  temperature  quartz  remains  indifferent,  but 
at  a  very  bright  heat  it  decomposes  the  salt,  forming  sodium 
silicate  and  chlorine.  There  are  works  in  operation  in  which, 
for  want  of  sulphur  in  the  ore,  the  chlorination  of  the  silver  is 
produced  partly  by  this  reaction  and  partly  by  the  chloridizing 
action  of  volatilized  salt.  It  requires  a  high  heat  and  a  large 
percentage  of  salt.  Quartz  is  the  most  desirable  gangue  in 
chloridizing  roasting. 

Carbonate  of  Lime  (Lime  Rock).  —  This  mineral,  which  occurs 
quite  frequently  as  gangue,  or  part  of  the  gangue,  acts  as  a  rule 
unfavorably  in  chloridizing  roasting.  It  takes  an  active  part  in 
the  process.  It  combines  with  the  sulphuric  acid  which  is  pro- 
duced by  the  combustion  of  the  metal  sulphides,  and  it  decom- 
poses also  the  metal  sulphates,  forming  sulphate  of  lime  and 
metal  oxides,  thus  preventing  them  from  acting  on  the  salt.  It 
decomposes  also  metal  chlorides,  forming  calcium  chloride  and 
metal  oxides.  Calcium  sulphate  is  indifferent  and  does  not  act 
on  the  salt.  If  there  is  more  carbonate  of  lime  in  the  ore  than 
can  be  converted  into  sulphate  and  chloride,  part  of  it  will  be 
found  in  the  roasted  ore  as  caustic  lime,  which  acts  decomposingly 
on  the  silver  chloride,  especially  so  in  the  subsequent  treatment 
for  extraction,  causing  a  poor  result.  If,  however,  there  are 
more  sulphides  in  the  ore  than  necessary  to  convert  the  lime  into 
sulphate  and  chloride,  usually  a  good  chlorination  of  the  silver 
can  be  obtained,  with  the  further  advantage  that  the  final  silver 
precipitate  will  be  very  rich  in  silver,  almost  free  from  base- 
metal  sulphides,  and  easily  convertible  into  metallic  silver  of  great 
fineness.  Therefore,  if  lime  is  present  in  the  ore  in  moderate 
quantities  it  is  beneficial  to  chloridizing  roasting.  The  loss  of 
silver  by  volatilization  will  be  found  moderate,  as  most  of  the 
volatile  chlorides  are  converted  by  the  lime  into  oxides,  which 
then  are  not  volatile  and  will  not  induce  silver  chloride  to  vola- 
tilize. 

Porphyry,  Clay,  Slate,  and  Other  Gangue  Minerals  Containing 
Alumina.  —  F.  Sustersic  made  the  very  interesting  observation 
that  under  certain  conditions  a  great  loss  of  silver  may  be 
caused  by  the  presence  of  alumina.  The  chlorine  acts  on  the 
alumina,  forming  aluminum  chloride,  which  is  extremely  volatile 
and  induces  the  silver  to  volatilize.  The  conditions  under  which 


THEORY  OF  CHLORIDIZING  ROASTING  9 

this  unfavorable  reaction  takes  place  were  not  ascertained.  As  a 
rule  the  gangue  minerals  named  in  this  paragraph  are  more  or 
less  indifferent;  and  do  not  exercise  a  bad  influence  in  chloridizing 
roasting. 

CLASSIFICATION  OF  ORES  IN  RELATION  TO 
CHLORIDIZING  ROASTING 

By  the  above-described  behavior  of  the  different  minerals 
in  chloridizing  roasting  it  is  apparent  that  chloridizing  roasting 
of  complex  silver  ores  is  undoubtedly  one  of  the  most  delicate  of 
metallurgical  operations.  The  treatment  has  to  be  modified  in 
accordance  with  the  character  of  the  ore,  and  the  character  of  an 
ore  in  relation  to  chloridizing  roasting  depends  on  the  nature  of 
the  different  sulphide  minerals  and  the  gangue  accompanying 
them.  The  sulphide  minerals  can  be  classified  as: 

(1)  Those,  like  iron  and  copper  pyrites,  gray  copper  ore,  silver 
copper  glance,  and  argentite,  which  form  in  roasting  sulphates 
which  act  on  the  sodium  chloride  and  liberate  the  chlorine. 

(2)  Those,  like  galena  and  zinc  blende,  which  form  sulphates 
remaining  indifferent  to  sodium  chloride. 

(3)  Antimonial    and   arsenical    silver  minerals,  which    form 
antimonates  and  arsenates  of  silver. 

The  gangue  either  remains  indifferent,  like  quartz  and  por- 
phyry, or  it  takes  an  active  part,  like  limestone,  and  minerals  con- 
taining magnesia. 

If  an  ore  consists  of  minerals  of  the  first  class,  together  with 
an  indifferent  gangue,  chloridizing  roasting  offers  no  difficulties 
nor  does  it  require  much  skill,  and  a  high  chlorination  can  be 
obtained  without  much  loss  of  silver  by  volatilization;  nor  does 
it  matter  whether  the  salt  is  added  to  the  charge  before  enter- 
ing the  furnace  or  after  it  has  been  subjected  to  a  partial  oxi- 
dizing roasting. 

The  process  of  chloridizing  roasting  becomes  more  difficult  if 
one  or  both  minerals  of  the  second  class  are  present  in  large 
quantities,  even  if  associated  with  an  indifferent  gangue.  The 
roasting  of  this  class  of  ore  is  elaborately  treated  in  Chapter  XI. 
With  such  ores  the  time  the  salt  is  added  becomes  very  impor- 
tant. If  added  before  the  charge  enters  the  furnace  a  very 
inferior  chlorination  is  obtained,  as  is  also  the  case  if  the  salt  is 
added  before  the  oxidizing  period  has  sufficiently  advanced,  or  if 


10  HYDROMETALLURGY  OF  SILVER 

it  is  added  when  the  period  has  too  far  advanced.     The  tempera- 
ture and  air  supply  require  much  attention. 

The  roasting  is  not  less  difficult  if  all  three  classes  are  repre- 
sented, especially  in  connection  with  a  gangue  like  limestone, 
which  takes  an  active  and  often  injurious  part  in  the  operation. 
This  class  of  ore  is  treated  elaborately  in  Chapter  XII. 


II 


CRUSHING  OF  THE  ORE 

THE  fineness  to  which  an  ore  has  to  be  reduced  in  order  to 
give  the  best  roasting  result  depends  on  the  chemical  and  physical 
character  of  the  material.  As  a  rule,  finely  pulverized  ore  roasts 
quickly  and  gives  a  better  result  than  a  coarser  material.  Ores 
which  decrepitate  when  charged  in  the  furnace,  or  ores  which 
during  the  combustion  of  the  sulphur  swell  and  disintegrate  like 
iron  pyrites,  can  be  crushed  rather  coarse  and  still  will  give  good 
chloridizing  results.  The  ore  of  Sombrerete,  Zacatecas,  Mexico, 
gave  good  roasting  results  if  crushed  through  a  screen  with  10, 
even  with  8,  meshes  to  the  linear  inch,  though  the  ore  contained 
much  zinc  blende  and  galena.  The  zinc  blende,  however,  was  of 
that  kind  which  decrepitates,  and  besides,  the  ore  was  crushed  in 
a  stamp  battery.  In  crushing  in  a  battery  the  larger  portion  of 
the  material  is  much  finer  than  the  size  of  the  meshes  calls  for. 
This  is  particularly  the  case  with  heavy  ores.  It  is  doubtful  if 
the  same  good  result  could  have  been  obtained  if  the  same  ore 
had  been  crushed  through  rolls,  because  rolls  produce  a  pulp 
much  more  uniform  in  size,  with  a  much  larger  percentage  corre- 
sponding in  size  with  the  size  of  the  meshes  of  the  screen. 

I  made  some  experiments  in  this  direction  with  ore  of  the 
San  Francisco  del  Oro  mine,  near  Santa  Barbara  and  Parral, 
Chihuahua,  Mexico.  The  ore  consists  principally  of  a  very  dense 
zinc  blende  and  finely  divided  galena.  The  zinc  blende  did  not 
decrepitate.  The  zinc  blende  and  the  galena  were  the  principal 
silver-bearing  minerals  of  the  ore. 

A  series  of  roasting  experiments  was  made  with  ore  crushed 
through  20-  and  through  40-mesh  screens.  The  ore  was  crushed 
in  a  stamp  battery.  It  was  found  that  the  ore  crushed  through 
20-mesh  required  a  much  longer  time  and  was  27  per  cent,  less 
chloridized  than  the  ore  crushed  through  the  40-mesh  screen. 
The  material  which  passes  through  a  battery  screen  of  certain 

11 


12 


HYDROMETALLURGY  OF  SILVER 


size  is  much  finer  than  the  size  of  the  meshes.  Heavy  ore  makes 
a  much  finer  pulp  through  the  same  screen  than  lighter  ore. 
The  pulp  of  the  Del  Oro  ore,  obtained  by  crushing  through 
battery  screens  No.  20  and  No.  40,  was  sifted  through  sieves  of 
different  fineness,  and  the  following  figures  obtained: 


BATTERY  PULP 
WHEN  SIFTED 
THROUGH  SIEVE 

CRUSHED 
THROUGH  SCREEN 
No.  20 

CRUSHED 
THROUGH  SCREEN 
No.  40 

CRUSHED 
THROUGH  SCREEN 
No.  20 

CRUSHED 
THROUGH  SCREEN 
No.  40 

PERCENTAGE  OF 
MATERIAL  PASS- 
ING THROUGH 
THE  SIEVE 

PERCENTAGE  OF 
MATERIAL  PASS- 
ING THROUGH 

THE  SIEVE 

PERCENTAGE  OF 
MATERIAL 
REMAINING  ON 
THE  SIEVE 

PERCENTAGE  OF 
MATERIAL 
REMAINING  ON 
THE  SIEVE 

No.  30 
No.  40 
No.  60 
No.  80 
No.  90 

93.8 
87.3 
78.8 
71.2 
67.1 

100 
100 
98.95 
93.80 
90.50 

6.2 
12.7 
20.2 
28.7 
32.9 

0.0 
0.0 
1.05 
6.20 
9.50 

These  figures  show  how  exceedingly  fine  a  heavy  ore  is  crushed 
in  a  battery,  even  through  a  screen  with  comparatively  coarse 
meshes.  Though  67.1  per  cent,  of  the  material  which  was  crushed 
through  screen  No.  20  was  finer  than  sieve  No.  90,  the  average 
chlorination  of  quite  a  number  of  compared  roastings  was  27  per 
cent,  less  than  that  of  ore  crushed  through  battery  screen  No.  40. 
This  indicates  how  essential  it  is  to  crush  such  ores  fine. 

It  is  frequently  argued  in  favor  of  coarse  crushing  that  coarser 
crushed  ore  permits  in  the  subsequent  lixiviation  a  free  percola- 
tion of  the  solution. 

While  to  a  certain  extent  coarsely  crushed  ore  permits  a 
somewhat  quicker  filtration,  the  increase  (if  extremes  are  avoided) 
is  slight  and  of  not  much  practical  value.  If  a  finely  crushed  ore 
filters  too  slow  for  an  extraction  by  filtration  it  will  filter  too 
slow  if  it  is  crushed  coarser,  because  in  crushing  always  a  certain 
amount  of  very  fine  powder  (slime)  is  formed,  no  matter  what 
kind  of  a  pulverizing  machine  is  used,  and  if  the  nature  of  the 
ore  is  such  as  not  to  undergo  much  of  a  physical  change  in  roast- 
ing, the  pulp  in  either  case  will  contain  sufficient  slimes  to  inter- 
fere with  a  free  percolation.  A  free  percolation  does  not  depend 
on  the  coarseness  of  the  pulp  nor  on  the  nature  of  the  gangue; 
it  depends  almost  entirely  on  the  nature  of  the  sulphides  and  on 
the  proportion  of  metal  sulphides  and  the  gangue.  Besides  the 
chemical  changes  which  an  ore  undergoes  during  chloridizing 


CRUSHING  OF  THE  ORE  13 

roasting,  a  change  of  its  physical  condition  also  takes  place. 
Lead  sulphate,  which  is  formed  in  roasting,  melts  easily  at  a 
roasting  temperature  and  is  absorbed  by  the  gangue  and  metal 
oxides.  The  same  is  the  case  with  cuprous  and  cupric  and  with 
ferrous  and  ferric  chlorides.  They  melt  even  below  red  heat  and 
also  penetrate  the  ore.  By  doing  so,  these  metal  salts  collect 
all  the  dusty  particles  or  slimes  of  the  gangue  and  metal  oxides 
into  small  porous  globules  and  flakes,  in  which  changed  condi- 
tion the  ore  permits  a  free  percolation.  This  is  the  cause  why  a 
chloridized  ore  filters  so  much  better  than  a  raw  ore,  and  if  the 
ore  contains  a  sufficient  amount  of  metal  sulphides  it  will  filter 
well  whether  crushed  very  fine  or  whether  it  is  crushed  coarser. 
If  finely  pulverized  the  conditions  for  the  chemical  reactions, 
however,  are  much  more  favorable. 

The  melting  of  the  metal  chlorides  and  lead  sulphate  and 
their  absorption  by  the  ore  causes  the  loosening  and  swelling  of 
the  charge,  making  it  what  is  called  "  woolly,"  during  the  chlorid- 
izing  period.  It  assumes  a  moist  appearance  and  can  be  stirred 
without  dusting,  and  does  not  evade  the  hoe  as  during  the  oxi- 
dizing period,  but  can  be  banked  and  collected  into  a  pile.  To 
maintain  this  condition  the  charge  has  to  be  agitated  from  time 
to  time,  otherwise  a  crust  will  be  formed  on  the  surface. 

If  for  want  of  a  sufficient  amount  of  sulphides  in  the  ore  the 
formed  chlorides  and  sulphates  are  insufficient  to  cause  this 
physical  change,  the  ore  will  remain  dusty,  run  like  water  on  the 
cooling  floor,  and  will  filter  very  slowly.  On  the  other  hand,  if 
the  amount  of  sulphides,  especially  lead  sulphide,  is  too  great 
in  proportion  to  the  gangue  and  metal  oxides,  the  latter  will 
get  so  saturated  that  they  cannot  maintain  their  loose  condition, 
and  form  lumps.  If  the  temperature  is  kept  moderate  these 
molten  chlorides  and  lead  sulphates  will  act  like  a  cement,  but  will 
not  go  into  chemical  combination  with  the  silica,  and  in  most 
cases  the  lumps  will  be  found  to  be  porous  and  soft  and  as  well 
roasted  as  the  finer  part.  But  if  the  heat  is  kept  too  strong 
silicates  will  be  formed  and  the  lumps  will  become  dense  and 
hard,  and  the  chlorine  will  be  unable  to  penetrate  them  and  act 
on  the  silver.  Particles  of  undecomposed  sulphides  will  be  en- 
closed in  them  and  cause  a  poor  extraction.  The  silver  can  be 
extracted  from  such  lumps  only  if  they  are  ground  and  reroasted 
with  steam,  by  which  hydrochloric  acid  is  formed,  which  acts  on 


14  HYDROMETALLURGY  OF  SILVER 

the  silicates.  Without  steam  only  a  small  percentage  of  the 
silver  can  be  extracted. 

Iron  sulphides  do  not  participate  so  much  in  changing  the 
physical  condition  of  the  ore  as  lead  or  copper  sulphides  do, 
because  the  chlorides  of  iron  easily  give  off  their  chlorine  and 
change  into  oxide,  which  then  acts  like  gangue.  The  most  effec- 
tive agent  is  the  lead  sulphide,  the  main  part  of  which  is  changed 
into  sulphate,  which  is  but  very  little  volatile  at  a  roasting  heat 
and  does  not  undergo  any  further  changes,  thus  much  improving 
the  filtering  quality  of  an  ore. 

In  working  the  refuse  dump  of  the  Cusihuiriachic  mine,  Chihua- 
hua, Mexico,  containing  from  25  to  30  oz.  silver  per  ton,  I  found 
that,  while  a  satisfactory  chlorination  of  the  silver  could  be 
obtained,  the  silver  could  not  profitably  be  extracted  on  account 
of  the  exceedingly  slow  filtration  caused  by  too  great  an  excess  of 
porphyry  gangue.  It  occurred  to  me  to  add  a  small  percentage 
of  galena,  and  the  effect  was  very  gratifying  —  the  ore  filtered 
well.  Later  the  slow  filtration  trouble  was  overcome  by  applying 
trough  lixiviation. 


Ill 

PERCENTAGE  OF  SALT  REQUIRED 

IF  all  the  chlorine  of  the  salt  could  be  transferred  to  the  silver 
only  an  insignificant  amount  of  salt  would  be  required,  but  as 
other  metals,  which  usually  are  present  in  much  larger  quantities 
than  silver,  are  also  chloridized,  a  correspondingly  large  percent- 
age of  salt  has  to  be  added  to  the  ore.  The  amount  to  be  added 
depends  on  the  nature  of  the  ore  and  has  to  be  ascertained  em- 
pirically in  each  individual  case.  It  is  best  to  commence  with  a 
high  percentage,  say  10  per  cent.,  of  salt,  and  to  reduce  the  salt  1 
per  cent,  in  each  succeeding  roasting  charge  until  3  per  cent,  is 
reached.  Ores  which  can  be  chloridized  with  less  than  3  per 
cent,  are  very  rare.  The  roasted  charges  are  tested  in  the  lab- 
oratory for  silver  chloride.  It  will  be  found  in  most  cases  that 
10  per  cent,  of  salt  does  not  produce  a  higher  chlorination  than 
6  or  5  per  cent.,  and  the  experimenter  will  decide  on  the  least 
amount  of  salt  which  produces  as  good  a  chlorination  as  the  next 
larger  amount,  and  will  adopt  that  percentage.  There  are 
instances,  however,  where  it  will  be  found  of  advantage  not  to 
produce  the  highest  possible  chlorination,  but  to  be  contented 
with  a  somewhat  inferior  extraction.  This  is  the  case  when  the 
cost  of  the  extra  amount  of  salt  exceeds  the  value  of  the  additional 
amount  of  silver  gained.  This  occurs  usually  in  treating  the 
lower  grade  ores  in  remote  localities,  where  the  price  of  salt  is 
high. 

Ores  containing  a  large  percentage  of  lead  and  zinc  require 
less  salt  than  ores  rich  in  iron  and  copper  sulphides,  because  the 
main  part  of  the  lead  is  converted  into  lead  sulphate,  which  re- 
mains indifferent  during  the  chloridizing  period  and  does  not 
consume  any  chlorine.  This  is  also  the  case  with  zinc,  which  is 
mostly  converted  into  zinc  sulphate  and  oxide,  which  remain 
indifferent.  Most  of  the  iron  and  copper,  however,  is  converted 
first  into  chlorides  before  they  change  into  oxides,  and  of  course 

15 


16  HYDROMETALLURGY  OF  SILVER 

these  are  heavy  consumers  of  chlorine,  and  the  ore  therefore  re- 
quires more  salt.  For  instance,  the  ore  of  the  San  Francisco  del 
Oro  mine  in  Mexico,  which  is  very  heavily  mineralized,  containing 
zinc  24.08  per  cent.,  lead  11.92  per  cent.,  iron  7  per  cent.,  copper 
0.5  per  cent.,  and  sulphur  21.35  per  cent.,  required  only  3J  to  4 
per  cent,  of  salt. 

An  excess  of  salt  does  not  improve  chlorination;  on  the  con- 
trary, in  many  instances  I  have  observed  that  the  chlorination 
already  gained  was  reduced  by  adding  more  salt.  For  this  and 
for  economical  reasons  an  excess,  therefore,  should  be  avoided, 
especially  as  undecomposed  salt  in  the  roasted  ore  is  not  advan- 
tageous in  the  subsequent  extraction. 

THE  PROPER  TIME  TO  ADD  THE  SALT 

The  generation  of  chlorine  in  the  furnace  does  not  commence 
until,  by  the  oxidation  of  the  sulphur,  metal  sulphates  have  formed, 
which  then  act  on  the  salt.  The  first  part  of  the  process,  there- 
fore, is  an  oxidizing  process,  whether  the  ore  contains  salt  or  not, 
and  in  this  respect  it  would  be  immaterial  at  what  time  the  salt 
were  added.  That  the  ore  sustains  a  heavier  loss  of  silver  by  vola- 
tilization if  the  salt  is  added  before  the  oxidizing  period  is  not 
conclusively  proved,  and  actually  there  is  no  reason  for  it. 
Chlorides  are  not  formed  until  the  sulphates  are  formed,  and 
therefore  the  presence  of  salt  cannot  cause  a  greater  volatiliza- 
tion of  the  silver.  An  ore  which  is  apt  to  lose  silver  on  account 
of  its  arsenic  and  antimony  sustains  the  larger  part  of  its  loss 
during  oxidizing  roasting. 

There  are  ores,  however,  which  cannot  be  chloridized  success- 
fully if  the  salt  is  added  to  the  ore  in  the  beginning.  This  is  the 
case  with  ores  which  contain  a  large  percentage  of  a  dense  argen- 
tiferous zinc  blende,  or  argentiferous  galena,  or  both,  as  the  prin- 
cipal silver-bearing  minerals  of  the  ore.  The  reason  why  such 
ores  have  to  be  first  subjected  to  an  oxidizing  roasting  before  the 
salt  is  added  is  the  following: 

Zinc  blende,  if  subjected  to  oxidizing  roasting,  changes  into 
zinc  oxide  and  zinc  sulphate,  while  sulphurous  acid  escapes. 
The  process  of  oxidizing  the  zinc  blende  progresses  but  slowly, 
especially  if  the  mineral  is  very  dense.  Iron  and  copper  sulphides, 
on  the  other  hand,  oxidize  easily  and  are  converted  into  sulphates 


PERCENTAGE  OF  SALT  REQUIRED          17 

long  before  this  is  the  case  with  the  zinc  and  lead  sulphides. 
Zinc  and  lead  sulphates  do  not  act  decomposingly  on  the  salt, 
while  iron  sulphate  does  so  energetically.  Now,  if  the  ore  and 
salt  are  charged  together  we  will  find  that  the  iron  sulphate,  as 
soon  as  it  is  formed,  will  act  on  the  salt,  producing  chlorine  and 
transforming  itself  into  chloride  and  oxide.  The  chlorides  of 
iron  are  volatile,  and  also  give  off  the  chlorine,  changing  into  oxide. 
While  this  process  is  going  on  the  zinc  and  lead  sulphides  are  only 
partly  oxidized,  and  as  the  chlorine  in  roasting  has  but  very  little 
effect  on  the  raw  zinc  blende  and  galena,  the  silver  contained 
therein  will  not  be  chloridized  by  the  time  the  generation  of 
chlorine  and  the  action  of  iron  chloride  has  ceased.  The  conse- 
quence is  a  very  inferior  roasting  result.  If,  however,  the  ore  is 
charged  into  the  furnace  without  salt  and  subjected  to  an  oxi- 
dizing roasting  until  the  zinc  and  lead  sulphides  are  oxidized,  •  or 
to  a  certain  extent  oxidized,  and  then  the  salt  is  added,  the  gen- 
erated chlorine  and  the  iron  chlorides  will  find  the  silver  in  a  state 
in  which  it  will  combine  with  the  chlorine.  Iron  sulphate  re- 
quires considerable  heat  to  be  decomposed  directly  into  oxide  and 
sulphuric  acid,  and  if  the  heat  during  oxidizing  roasting  is  kept 
low,  there  will  be  sufficient  iron  sulphate  in  the  charge  to  decom- 
pose the  salt,  and  a  quite  satisfactory  chlorination  of  the  silver 
will  be  effected.  To  this  class  of  ores  belong  those  of  the  San 
Francisco  del  Oro  mine,  Chihuahua,  Mexico,  and  of  Sombrerete, 
Zacatecas,  Mexico. 

Another  instance  of  great  difference  in  the  behavior  of  the  ore, 
whether  the  salt  was  added  to  the  ore  in  the  battery  or  during 
the  oxidizing  period  in  the  furnace,  I  experienced  during  my 
investigation  of  the  chloridizing  roasting  of  the  calcareous  arseni- 
cal silver  ore  at  Yedras,  Sinaloa,  Mexico.  The  gangue  of  this 
ore  consisted  of  silicious  limestone  and  calcspar,  while  the  ore 
proper  consisted  of  argentiferous  arsenical  pyrites,  a  moderate 
amount  of  fine-grained  black  zinc  blende,  arsenical  fahlerz,  and 
some  iron  pyrites.  When  the  roasting  was  done  in  the  Bruckner 
furnace  there  was  a  marked  difference  in  the  behavior  of  the  ore. 
When  the  salt  was  added  in  the  battery,  the  ore  swelled,  became 
woolly,  kept  on  one  side  of  the  revolving  furnace,  and  when 
discharged  did  not  dust  and  remained  in  a  pile  on  the  cooling 
floor.  When  the  salt  was  added  toward  the  end  of  the  oxidizing 
period,  the  ore  did  not  assume  the  moist  appearance  so  charac- 


18  HYDROMETALLURGY  OF  SILVER 

teristic  in  chloridizing  roasting,  but  remained  very  loose  and 
level  in  the  revolving  cylinder,  and  when  discharged  made  much 
dust  and  ran  on  the  cooling  floor  like  water.  The  percentage  of 
silver  chlorination  was  in  both  cases  about  the  same,  but  the  ore 
which  contained  the  salt  at  the  beginning  formed  a  very  large 
amount  of  hard  balls,  which  increased  in  size  as  the  roasting 
progressed.  They  consisted  of  concentric  layers  and  were  smooth 
and  hard.  They  were  well  chloridized,  but  the  silver  could  not 
be  extracted  unless  they  were  first  pulverized,  as  they  were  too 
dense  to  permit  the  solution  to  percolate  through  them. 

The  sulphureted  part  of  the  ore  had  no  tendency  to  form 
lumps,  as  numerous  experiments  with  concentrates  of  the  same 
ore  showed.  In  this  case  we  have  an  instance  in  which  the  time 
of  adding  the  salt  was  conditioned  by  the  nature  of  the  gangue 
(see  Chapter  XII). 

If  the  salt  is  added  later,  it  is  not  necessary  to  dry  and  pul- 
verize it;  in  fact  it  is  better  not  to  do  it.  It  saves  expense, 
and,  besides,  it  is  difficult  to  spread  it  uniformly  over  the  charge, 
and  in  places  where  more  salt  drops  it  is  apt  to  form  lumps.  The 
action  of  finely  pulverized  salt  commences  immediately  on  touch- 
ing the  ore,  and  in  doing  so  it  becomes  sticky,  which  makes  it 
difficult  to  divide  and  to  mix  it  evenly.  This  is  still  more  the 
case  in  a  Bruckner  furnace.  If  coarse  salt  is  added,  the  crystals, 
which  usually  are  of  the  size  of  beans  and  have  more  or  less  moist- 
ure, coming  in  contact  with  the  hot  ore  decrepitate  quite  rapidly. 
The  particles  fly  in  all  directions,  striking  the  roof  and  sides  and 
falling  back  to  the  ore.  When  decrepitation  ceases  the  salt 
will  be  found  much  more  evenly  scattered  over  the  charge  than 
this  can  be  done  by  a  shovel,  and  the  disintegrated  particles  are 
small  enough  for  the  purpose.  The  chemical  action  does  not 
commence  quite  as  soon  as  with  pulverized  salt,  and  a  much 
better  mixing  can  be  secured.  Of  course,  larger  lumps  of  crystals 
cemented  together,  or  pieces  of  salt  crust,  have  to  be  mashed 
first.  Salt  fuses  and  is  absorbed  by  the  ore,  thus  coming  in  inti- 
mate contact  with  the  sulphates.  A  Bruckner  furnace  should 
not  be  set  to  revolve  until  decrepitation  ceases. 

To  extend  the  oxidizing  roasting  to  such  a  degree  as  to  produce 
a  "dead  roast,"  that  is,  to  convert  all  the  convertible  sulphates 
into  oxides  and  then  to  produce  the  chlorination  by  an  addition 
of  a  mixture  of  calcined  copperas  (ferrous  sulphate)  and  salt,  is 


PERCENTAGE  OF  SALT  REQUIRED  19 

by  far  too  slow  and  expensive  a  method  to  be  adopted  in  prac- 
tice. 

To  add  the  salt  during  crushing  produces  a  very  uniform 
mixture  of  ore  and  salt  and  simplifies  operations  in  roasting,  for 
which  reason  it  is  preferable  to  do  so  if  the  nature  of  the  ore 
permits  it;  still  it  is  not  frequently  done  in  practice,  unless  the 
construction  of  the  furnace  demands  it,  because  by  crushing  ore 
and  salt  together  the  crushing  capacity  of  the  machinery  is 
reduced  by  the  amount  of  salt  added,  and  even  more  if  the  latter 
is  not  previously  very  well  dried. 


IV 


LOSS  OF  SILVER  BY  VOLATILIZATION 

SILVER  chloride  as  such  is  not  volatile,  but  if  influenced  by 
the  volatilization  of  other  chlorides  it  becomes  volatile.  A  high 
heat,  therefore,  indirectly  causes  a  larger  loss  of  silver  by  the 
expulsion  of  larger  quantities  of  volatile  chlorides.  Other  condi- 
tions being  equal  we  shall  always  find  the  loss  of  silver  to  be  in 
direct  proportion  to  the  chemical  loss  in  weight  an  ore  sustains. 
In  other  words,  the  charge  of  the  same  ore  that  during  roasting 
sustains  the  least  chemical  loss  in  weight  sustains  also  the  least 
loss  of  silver  by  volatilization.  The  term  "chemical  loss  in 
weight"  is  used  in  distinction  to  the  loss  an  ore  sustains  during 
roasting  by  dusting,  which  is  a  mechanical  loss. 

The  logical  consequence  of  the  above  facts  is  that  the  operator, 
while  he  endeavors  to  obtain  a  high  silver  chlorination,  should  be 
at  the  same  time  careful  to  expel  as  little  as  possible  of  the  volatile 
chlorides.  He  will  be  greatly  assisted  in  this  endeavor  by  keeping 
the  ore  in  a  thick  layer,  and  by  using  low  heat  and  plenty  of  air. 
If  a  small  charge  is  thinly  spread  over  a  large  hearth  more  volatile 
chlorides  will  be  expelled,  and  the  ore  will  lose  more  in  weight 
and  in  silver  than  when  a  larger  charge  is  roasted  in  the  same 
furnace.  This  is  the  reason  why,  as  a  rule,  the  loss  in  weight 
and  in  silver  in  a  large  Bruckner  furnace,  in  which  the  ore  lies 
two  feet  thick,  is  less  than  in  a  reverberatory,  and  why  small 
samples  roasted  on  a  roasting  dish  in  the  muffle  show  so  much 
greater  loss  of  silver  than  the  same  ore  does  when  roasted  on  a 
large  scale  in  the  furnace. 

It  will  be  found  that  ore  roasted  at  a  low  heat  with  sufficient 
air  will  lose  less  in  weight,  because  a  large  part  of  the  volatile 
chlorides,  which  at  a  higher  heat  would  be  expelled,  will  then 
remain  in  the  ore.  For  amalgamation  it  is  desirable,  in  fact 
necessary,  to  expel  the  volatile  chlorides  as  much  as  possible, 
because  they  take  an  active  part  in  amalgamation  and  make  the 

20 


LOSS   OF  SILVER   BY  VOLATILIZATION  21 

quicksilver  smeary  and  inactive,  causing  a  poor  silver  extraction 
and  a  very  low-grade  bullion.  These  chlorides,  however,  do  not 
seriously  interfere  in  the  lixiviation  process;  in  fact,  it  is  one  of 
the  principal  advantages  of  lixiviation  over  amalgamation  that 
in  it  the  volatile  chlorides  do  not  need  to  be  expelled,  and  therefore 
the  roasting  of  most  ores,  even  those  rich  in  arsenic  and  antimony, 
can  be  conducted  with  a  very  small  loss  of  silver  by  volatilization. 
In  metallurgical  books  we  always  find  the  great  loss  of  silver 
pointed  out  as  an  objection  to  all  processes  which  require  chlori- 
dizing  roasting  of  the  ore.  Formerly  chloridizing  roasting  was 
principally  used  and  studied  in  relation  to  amalgamation;  little 
or  no  attention  was  paid  to  roasting  for  lixiviation,  or  to  the  fact 
that  this  process  allowed  a  modification  of  roasting  by  which  its 
objectionable  features  could  be  obviated. 

I  made  chloridizing  roasting  the  subject  of  special  study,  and 
found  that  it  could  be  conducted  with  just  as  little  loss  of  silver 
as  oxidizing  roasting,  if  care  was  taken  to  expel  as  little  as  pos- 
sible of  the  volatile  chlorides.  The  chemical  reaction  between 
salt  and  the  sulphates  takes  place  at  a  very  low  heat,  in  fact  at 
a  lower  heat  than  is  generally  believed,  while  on  the  other  hand 
it  takes  quite  a  high  heat  to  expel  thoroughly  the  volatile  chlo- 
rides; therefore,  in  roasting  for  lixiviation  the  temperature  can  be 
kept  as  low  as  the  nature  of  the  ore  permits  during  oxidizing,  and 
lower  still  during  chloridizing,  and  yet  have  the  ore  well  prepared 
for  the  subsequent  extraction  of  the  silver.  During  chloridizing 
the  ore  ought  to  be  kept  in  a  thick  layer  and  stirred  only  at 
intervals  to  diminish  the  volatilization  of  the  chlorides. 

In  the;  old  method  of  chloridizing  roasting  the  aim  was  to  free 
the  ore  by  heat  from  metal  chlorides  that  are  objectionable  for  the 
subsequent  extraction  of  the  silver,  while  in  the  new  method  the 
aim  is  to  retain  in  the  ore  as  much  of  the  chlorides  as  possible 
and  to  remove  them  by  leaching  with  water  previous  to  the  ex- 
traction of  the  silver.  If  we  take  into  consideration  the  fact  that 
the  otherwise  not  volatile  silver  chloride  becomes  volatile  by  the 
volatilization  of  other  metal  chlorides,  it  is  quite  logical  that  the 
volatilization  of  the  silver  will  be  greatly  reduced  by  the  modified 
method. 

In  roasting  the  calcareous  arsenical  silver  ore  at  Yedras, 
Sinaloa,  Mexico,  by  the  modified  method,  with  plenty  of  air,  the 
loss  in  weight  was  only  3.5  per  cent,  and  the  loss  in  silver  by 


22  HYDROMETALLURGY  OF  SILVER 

volatilization  only  1.8  per  cent.,  while  if  roasted  by  the  old  method 
the  loss  in  weight  was  found  to  be  from  7  to  13  per  cent.,  while  the 
loss  of  silver  was  15  to  25  per  cent,  and  more. 

METHOD  OF  ASCERTAINING  THE  Loss  OF  SILVER  BY 
VOLATILIZATION 

In  order  to  roast  skilfully  it  is  of  great  importance  to  ascer- 
tain frequently  the  loss  of  silver  by  volatilization,  but  to  do 
this  it  is  necessary  to  know  the  loss  in  weight  the  ore  sustains. 
This,  however,  is  accompanied  with  great  difficulty  if  it  is  done 
in  the  old  way  by  actual  weighing  of  the  charge  before  and  after 
roasting,  necessitating  the  careful  cleaning  of  the  furnace  and 
the  dust-chambers  before  and  after  the  process.  In  many  cases 
this  is  not  possible  without  seriously  interfering  with  the  regular 
work,  and  at  all  events  it  is  accompanied  with  so  much  trouble 
and  expense  that  if  the  shrinkage  in  weight  is  once  ascertained, 
this  figure  is  used  in  all  subsequent  calculations,  though  the 
conditions  under  which  the  roasting  is  performed,  such  as  heat 
and  draft  or  the  character  of  the  ore,  may  have  changed.  That 
such  figures  are  not  very  reliable  will  be  readily  understood,  but 
still  more  incorrect  is  the  method  some  adopt  of  roasting  10  or 
20  grams  in  the  muffle  and  then  taking  the  difference  in  weight 
before  and  after  roasting  as  the  loss  in  weight  the  ore  sustains  in 
roasting;  by  this  means  the  loss  of  that  particular  sample  in  the 
muffle  is  ascertained,  but  not  the  loss  the  ore  would  lose  in  the 
furnace.  Just  as  incorrect  is  the  practice  of  roasting  10  grams  in 
the  muffle,  of  using  the  roasted  10  grams  for  an  assay,  and  of 
comparing  the  assay  value  per  ton  with  the  assay  value  per  ton 
of  the  raw  ore.  This  gives  us  only  the  amount  of  silver  this 
particular  sample  lost  by  volatilization,  but  it  gives  no  informa- 
tion as  to  how  much  the  ore  loses  if  roasted  in  the  furnace,  because 
the  conditions  under  which  the  roasting  in  the  two  cases  takes 
place  are  very  different  with  regard  to  temperature,  draft,  time, 
and  thickness  of  the  layer. 

To  conduct  the  roasting  properly  it  is  not  of  great  importance 
to  know  how  much  the  ore  loses  by  dusting,  for  this  is  merely  a 
mechanical  loss,  and  the  fine  ore  particles  carried  away  by  the 
draft  are  easily  collected  in  dust-chambers.  The  loss  due  to  the 
volatilization  of  the  chlorides  is  the  serious  one.  These  fumes 


LOSS  OF  SILVER  BY   VOLATILIZATION  23 

are  often  richer  in  silver  than  the  ore,  are  difficult  to  collect,  and 
easily  escape.  We  have,  therefore,  to  find  how  much  the  ore 
in  the  furnace  loses  in  weight  by  volatilization  in  order  to  ob- 
tain a  correct  basis  for  a  calculation  of  the  loss  of  silver  in  roast- 
ing. 

I  adopted  the  following  method,  which  gives  sufficiently  correct 
results  for  practical  purposes,  can  be  performed  in  the  assay 
office  in  a  few  hours,  and  is  at  all  events  more  correct  than  if  the 
loss  in  weight  of  the  ore  is  ascertained  by  actual  weighing  of  the 
charge  and  flue-dust. 

Ten  grams  of  the  raw  pulp,  containing  the  same  percentage 
of  salt  as  the  ore  in  the  furnace,  are  placed  in  a  roasting  dish 
and  roasted  in  the  muffle  for  half  an  hour  or  an  hour;  then  the 
sample  is  removed  from  the  muffle,  allowed  to  cool,  weighed, 
returned  to  the  muffle,  roasted  again  for  half  an  hour,  and  then 
weighed  again.  This  is  repeated  -until  two  weighings  are  alike, 
or  until  in  the  last  half-hour  the  ore  does  not  lose  more  than 
2  or  3  mg.;  then  the  difference  between  the  original  weight  and 
that  of  the  last  weighing,  expressed  in  percentage,  gives  the 
highest  possible  loss  the  raw  ore  can  suffer. 

Ten  grams  of  a  sample  of  roasted  ore,  corresponding  with  the 
sample  of  raw  pulp,  are  placed  in  a  roasting  dish,  and  also  roasted 
in  the  muffle  until  two  weighings  agree,  or  the  difference  between 
two  consecutive  weighings  is  not  more  than  2  or  3  mg.  The  dif- 
ference between  the  first  weighing  (10  grams)  and  the  last,  ex- 
pressed in  percentage,  gives  the  weight  which  the  roasted  ore  is 
still  capable  of  losing  if  subjected  to  prolonged  roasting.  If  we 
deduct,  therefore,  the  capable  loss  from  the  highest  possible  loss, 
we  obtain  in  percentage  the  loss  in  weight  the  ore  has  suffered 
during  roasting  in  the  furnace  by  volatilization. 

In  the  following,  the  weighings  are  given  of  one  of  the  tests 
which  I  made  with  ore  roasted  in  Bruckner  cylinders  at  Yedras, 
Mexico: 

RAW  ORE,  CONTAINING  7  PER  CENT.  SALT 

Original  weight 10  grams. 

After  1  hour  roasting  in  the  muffle 9.35   " 

After  J  hour  more  roasting  in  the  muffle 9.23   " 

After  |  hour  more  roasting  in  the  muffle 9.21    " 

Ten  grams  —  9.21  grams  =  0.79  grams  =  7.9  per  cent,  highest 
possible  loss  in  weight. 


24 


HYDROMETALLURGY  OF  SILVER 


ROASTED  ORE 

Original  weight  •••••• 10  grams. 

After  1  hour  roasting  in  the  muffle 9.65   " 

After  ^  hour  more  roasting  in  the  muffle 9.51   " 

After  \  hour  more  roasting  in  the  muffle 9.50  " 

Ten  grams  —  9.5  grams  =  0.5  grams  =  5  per  cent,  loss,  which 
the  ore,  roasted  in  the  furnace,  was  still  capable  of  sustaining 
by  dead  roasting. 

Highest  possible  loss  of  raw  ore 7.9  per  cent. 

Capable  loss  of  roasted  ore 5.0         " 

Actual  loss  in  weight  in  the  furnace 2.9  per  cent. 

The  gangue  of  the  Yedras  ore  is  limestone.  Agreeing  weights, 
however,  are  more  quickly  obtained  if  ores  have  quartz  gangue, 
while  ores  containing  considerable  manganese  take  a  longer  time, 
and  require  more  patience.  It  is  advisable  to  pulverize  the  ore 
carefully  once  or  twice  in  a  porcelain  mortar  during  the  test,  in 
order  to  break  up  small  lumps  which  have  formed. 

As  this  test  is  so  quickly  and  easily  done,  it  gives  the  metal- 
lurgist the  means  of  ascertaining  the  most  favorable  tempera- 
ture and  proper  time,  and  of  controlling  the  work  of  the  man 
in  charge  of  the  furnace.  The  mere  difference  of  assay  value 
between  raw  and  roasted  ore  is  no  guide,  as  can  be  seen  in  the 
following  table,  in  which  the  results  of  a  few  tests  are  given, 
which  I  made  with  the  ore  of  the  Cusihuiriachic  Silver  Mining 
Company,  Chihuahua,  Mexico: 


ASSAY 
VALUE  OF 
RAW  ORE 
PER  TON 
OUNCES 

ASSAY 
VALUE  OF 
ROASTED 
ORE 
PER  TON 
OUNCES 

HIGHEST 
POSSIBLE 
Loss  OF 
RAW  ORE 
PER  CENT. 

CAPABLE 
Loss  OF 
ROASTED 
ORE 
PER  CENT. 

ACTUAL 
Loss 
IN  WEIGHT 
SUSTAINED 
IN  ROASTING 
PER  CENT. 

Loss  OF 
SILVER  BY 
VOLATILI- 
ZATION 
PER  CENT. 

REMARKS 

46.0 

43.6 

7.0 

6.0 

1.0 

6.2 

41.8 

41.6 

7.3 

5.5 

1.8 

2.2 

43.2 

41.0 

7.5 

5.3 

2.2 

7.2 

Roasted  in  "C.'s"  shift 

43.2 

41.6 

7.5 

6.2 

1.3 

4.9 

Roasted  in  "L.'s"  shift 

51.0 

50.0 

7.8 

6.6 

1.2 

3.1 

The  roasting  was  done  in  Howell  furnaces.  Each  of  the 
above  tests  was  made  with  average  samples  of  a  whole  day's 
roasting.  The  third  and  fourth  sample,  however,  represent  the 
ore  roasted  in  one  day,  one  roasted  in  C.'s  shift  (night)  and  the 
other  in  L.'s  shift  (day).  Corresponding  samples  were  taken  of 
the  raw  ore.  The  salt  was  added  to  the  ore  in  the  battery.  The 


LOSS  OF  SILVER  BY  VOLATILIZATION  25 

assay  showed  that  the  value  of  the  raw  ore  was  in  both  shifts 
the  same,  while  the  assay  value  of  the  roasted  ore  of  both  men 
was  nearly  the  same,  L.'s  assaying  only  0.6  oz.  more.  Judging 
by  the  assays,  we  are  apt  to  think  that  the  work  of  both  men 
was  nearly  alike,  but  by  referring  to  the  column  showing  the 
percentage  of  silver  volatilization,  we  find  C.  lost  7.2  per  cent, 
while  L.  only  4.9  per  cent,  silver.  These  figures  also  tell  the  cause 
why  C.  lost  more  silver.  The  ore  roasted  by  him  showed  a  capable 
loss  in  weight  of  5.3  per  cent,  while  that  roasted  by  L.  showed 
a  capable  loss  in  weight  of  6.2  per  cent.  C.  therefore  roasted  at 
too  high  a  heat,  expelling  unnecessarily  more  volatile  chlorides, 
and  by  doing  so  increased  the  loss  of  silver. 


V 

METHODS  OF  ROASTING 
CHLORIDIZING  SELF-ROASTING 

THIS  mode  of  roasting  can  only  be  successfully  performed  with 
highly  sulphureted  ore  and  in  a  furnace  the  construction  of 
which  permits  the  roasting  of  large  charges,  like  the  Bruckner 
type  of  furnaces.  We  have  seen  above  that  in  roasting  for  the 
process  of  lixiviation  with  sodium  hyposulphite  it  is  not  necessary 
to  expel  the  metal  chlorides  by  increasing  the  heat  to  bright  red 
toward  the  end  of  the  chloridizing  period,  but  that,  on  the  con- 
trary, the  roasting  should  be  conducted  at  a  low  heat  to  the  very 
end,  to  retain  in  the  roasted  ore  as  much  of  the  metal  chlorides 
as  possible  in  order  to  reduce  the  loss  of  silver  by  volatilization. 
Reflecting  on  this  principle,  it  occurred  to  me,  while  roast- 
ing heavily  sulphureted  ore  in  a  Bruckner  furnace,  that  the 
charge  if  once  ignited  may,  by  the  oxidation  of  the  sulphides, 
produce  and  keep  in  store  sufficient  heat  to  finish  the  chloridiz- 
ing part  of  the  process  without  applying  any  additional  heat. 
Experiments  showed  that  this  could  be  successfully  done,  and 
that  not  only  was  50  per  cent,  of  the  fuel  saved,  but  that,  while 
the  chlorination  of  the  silver  was  5  per  cent,  higher,  the  loss  of 
silver  by  volatilization  was  materially  less  than  by  applying  a 
second  fire.  It  was  possible  in  this  way  to  roast  10.6  tons  of  ore 
with  one  cord  of  wood. 

The  charges  should  not  be  smaller  than  4J  to  5  tons,  otherwise 
the  heat  stored  in  the  ore  will  die  out  before  the  roasting  is  finished. 
When  the  furnace  is  charged,  a  strong  fire  is  kept  up  until  the  ore 
has  fairly  started  to  roast;  then  the  fire  is  allowed  to  go  out,  or 
if  necessary  pulled  out,  and  the  fire-door  left  open  to  allow  a 
sufficient  supply  of  air  to  pass  through  the  furnace.  The  heat 
gradually  increases  though  the  fire  is  out.  The  charge  maintains 
nearly  a  horizontal  position.  In  due  time  the  ore  loses  its  bright- 
ness, increases  in  volume,  and  begins  to  assume  a  more  erect 

26 


METHODS  OF  ROASTING  27 

position,  leaning  against  that  side  which  moves  upward.  The 
chloridizing  period  has  commenced.  While  during  oxidizing  the 
ore  looks  bright  and  the  furnace  lining  dark,  just  the  reverse 
can  be  observed  during  chloridizing:  the  surface  of  the  ore  looks 
dark  while  the  lining,  emerging  from  the  ore,  looks  red.  Of 
course,  that  part  of  the  ore  which  is  brought  up  by  the  motion 
of  the  furnace  is  also  red,  but  it  quickly  darkens. 

It  will  be  found  that  the  chlorination  of  the  silver  is  finished 
before  the  red  heat  of  the  charge  has  entirely  died  out,  and  this 
is  the  proper  time  to  discharge  the  furnace,  in  the  first  place  to 
avoid  loss  of  time,  and  secondly  to  avoid  dusting.  A  chloridized 
ore  when  still  red  does  not  dust  much  in  discharging,  while  when 
it  gets  completely  dark,  but  is  still  hot,  it  dusts  considerably  more 
than  if  handled  when  quite  cool. 

If  the  ore  is  rich  in  sulphides,  the  salt  can  be  added,  if  required, 
during  the  oxidizing  period,  but  this  ought  to  be  done  quickly  in 
order  not  to  cool  the  furnace  too  much. 

I  adopted  the  term  "chloridizing  self-roasting"  for  this  mode 
of  roasting  because,  after  the  ore  is  ignited  and  the  fire  is  removed, 
it  passes  through  the  oxidizing  and  through  the  chloridizing  period 
without  requiring  any  further  attention.  One  man  can  attend 
to  quite  a  number  of  furnaces. 

The  ore  thus  roasted  is  roasted  at  the  lowest  possible  temper- 
ature. 

CHLORIDIZING  HEAP-ROASTING 

If  silver  ore  which  has  been  subjected  to  chloridizing  roast- 
ing is  left  in  a  pile  when  discharged  from  the  furnace,  it  will 
retain  a  dark -red  heat  for  many  hours,  during  which  time  the 
process  of  chlorination  continues.  I  found  that,  if  the  chlorina- 
tion of  the  silver  is  accomplished  in  the  furnace  up  to  85  or 
90  per  cent.,  the  increase  in  chlorination  amounts  to  respectively 
2  and  1  per  cent.,  and  that  this  increase  takes  place  principally 
during  the  first  two  or  three  hours.  By  extending  the  time  only 
an  insignificant  increase  takes  place.  This,  however,  is  different 
if  the  chlorination  in  the  furnace  be  less  advanced  at  the  time 
of  discharge.  In  such  a  case  a  large  increase  in  chlorination 
takes  place  on  the  cooling  floor.  C.  A.  Stetefeldt  made  the  inter- 
esting and  valuable  observation  that  even  in  a  very  poorly  roasted 
ore  the  chlorination  of  the  silver  can  be  brought  up  to  a  high 


28  HYDROMETALLURGY  OF  SILVER 

percentage  if  the  ore  is  left  in  a  pile  on  the  cooling  floor.  In 
roasting  the  ore  of  the  Lexington  mine  he  found  it  to  be  of 
such  a  nature  that  a  silver  chlorination  of  only  about  47  per 
cent,  could  be  obtained  in  the  shaft  of  the  Stetefeldt  furnace. 
This  partially  roasted  ore  was  piled  -on  the  cooling  floor  while  his 
roasting  experiments  were  going  on.  An  examination  of  the 
roasted  ore  after  twelve  hours,  however,  showed  an  increase  in 
chlorination  of  from  47  to  90  per  cent.  The  ore  was  too  heavily 
charged  with  sulphide  to  be  suitable  for  a  complete  roasting  in 
this  furnace,  and  only  a  partial  oxidation  took  place,  but  when 
piled  in  a  heap  the  oxidation  continued,  forming  sulphates  which, 
acting  on  the  salt,  produced  the  chlorination.  The  temperature 
produced  by  the  slow  oxidation  was  sufficiently  high  for  the 
chemical  reaction.  This  observation  may  lead  to  the  adoption 
in  practice  of  a  new  method  of  chloridizing  roasting,  which  we 
properly  may  call  "chloridizing  heap-roasting." 

It  is  apparent  that,  if  a  chloridizing  roasting  could  be  per- 
formed just  by  exposing  the  ore  to  a  short  roasting  in  the  furnace, 
and  then  leaving  it  to  itself  in  a  pile  outside  the  furnace  until 
cool,  the  advantages  gained  would  be  great,  metallurgically  as 
well  as  financially.  This  method,  however,  is  only  applicable  to 
ores  not  too  heavily  charged  with  zinc  blende  and  galena,  as  I 
once  had  the  opportunity  to  convince  myself.  When  experi- 
menting with  the  heavy  zinc-lead  ore  of  the  San  Francisco  del  Oro 
mine,  Chihuahua,  Mexico,  containing  24  to  25  per  cent,  zinc, 
11.9  per  cent,  lead,  7  iron,  and  21  sulphur,  I  also  tried  the  Stete- 
feldt furnace.  This  was  done  more  to  obtain  positive  figures  and 
a  complete  record  of  my  investigation  than  in  expectation  of 
obtaining  satisfactory  results.  The  ore  when  removed  from  the 
shaft  of  the  furnace  emitted  large  volumes  of  sulphurous  acid  gas. 
No  chlorine  could  be  detected  and  the  chlorination  obtained  did 
not  exceed  15  to  16  per  cent.  The  ore  was  piled  on  the  cooling 
floor.  There  it  continued  to  roast,  emitting  sulphurous  acid  fumes 
for  several  days,  until  it  finally  cooled  without  showing  a  per- 
ceptible increase  in  chlorination. 

Heap-roasting  was  tried  again  by  me,  with  the  ore  of  Som- 
brerete,  Zacatecas,  Mexico,  which  contained  8.9  per  cent,  zinc, 
9.5  per  cent,  lead,  16.8  per  cent,  iron,  and  26.4  per  cent,  sulphur. 
Though  this  ore  is  considerably  lighter  in  zinc  and  lead  than  the 
ore  of  the  San  Francisco  del  Oro  mine,  it  was  still  too  heavy  to 


METHODS   OF  ROASTING  29 

be  tried  with  the  means  of  a  Stetefeldt  furnace.  In  order  to 
promise  success  the  oxidation  of  the  metal  sulphides  had  to  be 
brought  to  a  more  advanced  state  than  can  be  done  in  a  Stetefeldt 
furnace,  especially  as  the  Sombrerete  ore  also  required  an  oxidiz- 
ing roasting  to  a  certain  state  before  adding  the  salt.  The 
experiment  was  made  with  the  aid  of  reverberatory  furnaces. 
In  three  adjoining  reverberatory  furnaces  three  charges  of  one 
ton  each  were  oxidized  until  the  color  of  the  ore  commenced  to 
change  to  brown,  but  still  contained  many  black  particles,  and 
still  smelled  quite  strongly  of  sulphur.  Then  6  per  cent,  of  salt 
was  scattered  over  the  surface  of  the  ore.  Immediately  after 
adding  the  salt  the  three  furnaces  were  discharged  simultaneously, 
and  the  hot  ore  of  the  three  charges  was  piled  into  one  heap  in 
the  yard  outside  the  building  and  left  there  to  chloridize.  After 
lying  for  fourteen  hours,  it  was  found,  by  inserting  the  sampling 
rod,  that  the  ore  inside  the  pile  was  still  red  hot,  and  that  the 
fumes  of  the  sample  still  smelled  strongly  of  sulphurous  acid. 
The  color  had,  to  a  great  extent,  changed  from  brown  to  red. 
A  test  of  the  sample  showed  that  only  12.6  per  cent,  of  the  silver 
was  chloridized.  After  twenty-three  hours  it  was  found  that  the 
temperature  inside  the  heap  was  considerably  lower,  but  still 
high  enough  for  the  generation  of  chlorine.  A  distinct  odor  of 
chlorine  was  emitted  from  the  sample,  but  none  of  sulphur.  The 
chlorination  of  the  silver  had  increased  to  74.2  per  cent.  After 
thirty-eight  hours  the  ore  had  cooled  down  to  an  extent  that  no 
more  chemical  reaction  could  take  place.  The  heap  was  spread 
out  and  sampled.  The  color  of  the  ore  was  as  red  as  that  of 
charges  finished  in  the  furnace.  The  chlorination  of  the  silver 
was  found  to  have  increased  to  85  per  cent.  There  is  no  reason 
why  the  chlorination  could  not  have  been  raised  to  90  or  95  per 
cent,  and  higher,  if  the  proper  temperature  could  have  been 
maintained  longer,  but  the  heap  being  so  small,  containing  only 
three  tons,  it  lost  its  heat  before  the  chlorination  was  finished. 
Only  during  the  time  of  dumping  the  hot  ore  on  a  pile  could 
the  fumes  be  seen.  As  soon  as  the  pile  was  completed  visible 
fumes  ceased  to  emanate.  A  strong  odor  of  sulphurous  acid 
could  be  observed  for  quite  a  number  of  hours,  indicating  that 
oxidation  was  still  continuing,  but  no  fumes  could  be  seen. 
When  the  chloridizing  period  commenced  the  odor  of  sulphurous 
acid  ceased,  but  no  odor  of  chlorine  could  be  noticed  in  its  place, 


30  HYDROMETALLURGY  OF  SILVER 

nor  did  any  visible  fumes  emanate  from  the  pile.  But  from 
a  sample  taken  from  the  inside  of  the  pile  light  fumes  could  be 
observed,  accompanied  by  an  odor  of  chlorine,  indicating  that 
no  volatile  chlorides  were  emitted  from  the  pile,  and  that  the 
generated  chlorine  went  into  combination  with  the  metals  of  the 
ore.  The  ore  being  undisturbed  and  in  a  thick  layer,  an  excellent 
opportunity  existed  for  this  chemical  reaction.  In  roasting  in  a 
reverberatory  furnace  it  can  plainly  be  observed  that  the  ore 
on  the  hearths,  even  on  the  chloridizing  hearth,  will  not  emit 
much  visible  fume,  but  as  soon  as  the  ore  is  disturbed  by  the 
movements  of  the  rake  heavy  fumes  will  be  emitted.  Now. 
these  emanating  volatile  fumes  are  the  very  cause  of  the  volatili- 
zation of  the  silver.  It  is  therefore  apparent  that,  if  the  crea- 
tion of  such  volatile  metal  chlorides  can  be  avoided,  the  loss  of 
silver  will  be  reduced  to  the  minimum  —  that  is,  to  the  loss  which 
will  occur  during  oxidizing  roasting,  and  which,  in  most  cases, 
is  very  small. 

On  examination  of  the  ore  roasted  in  that  experimental  heap, 
it  was  found  that  much  more  metal  subchlorides  than  chlorides 
were  formed  as  compared  with  roasting  in  the  furnace.  As  a 
much  better  utilization  of  the  chlorine  takes  place  if  the  ore  is 
in  a  heap  and  left  undisturbed  than  when  spread  over  a  hearth 
in  a  comparatively  thin  layer,  it  is  to  be  expected  that  roast- 
ing in  heaps  will  require  less  salt.  This  agrees  with  observations 
I  have  made  by  roasting  the  same  ore  in  a  Bruckner  and  in  a 
reverberatory  furnace.  In  the  Bruckner  furnace  less  salt  was 
required  while  a  higher  chlorination  was  obtained,  together 
with  a  smaller  loss  of  silver  by  volatilization.  In  the  Bruckner 
furnace  the  ore  is  in  a  much  thicker  layer  than  in  the  reverbera- 
tory, which  causes  the  better  results. 

The  experiment  at  Sombrerete  was  made  under  very  un- 
favorable conditions.  The  heap  was  too  small,  containing  only 
three  tons,  and  was  exposed  from  all  sides  to  the  cooling  action 
of  the  air,  so  that  the  chemical  reaction  ceased  before  the  chlori- 
nation was  completed.  Notwithstanding  this,  the  results  ob- 
tained showed  that  85  per  cent,  of  the  silver  was  chloridized,  and 
if  we  take  into  consideration  the  increased  furnace  capacity,  the 
reduction  in  the  consumption  of  fuel  and  salt,  this  method  surely 
offers  sufficient  advantages  to  justify  further  investigations  and 
experiments.  To  maintain  favorable  conditions  the  hot  ore 


METHODS   OF  ROASTING  31 

should  be  dumped  into  bins  made  of  bricks  or  stone  masonry, 
holding  30  to  40  tons.  The  number  of  these  bins  will  depend  on 
the  roasting  capacity  of  the  works  and  on  the  time  a  heap  will 
require  to  complete  the  roasting.  The  bins  which  have  to  be 
placed  on  the  cooling  floor  should  be  open  on  the  side  toward 
the  cooling  floor,  or  provided  with  a  good  sized  iron  door,  to  per- 
mit free  access,  because  chloridized  ore  as  a  rule  does  not  run 
and  has  to  be  poked  down.  The  top  of  the  bins  should  be  pro- 
vided with  hoods,  to  take  off  the  sulphur  gas  which  will  emanate 
from  the  ore  for  some  time. 

Chloridizing  heap-roasting  may  prove  to  be  the  most  rational 
mode  of  chloridizing  roasting.  A  higher  percentage  of  silver 
will  be  chloridized  with  less  loss  of  silver  and  at  a  smaller  cost 
than  if  the  roasting  is  finished  in  the  furnace,  no  matter  what 
type  of  roasting  furnace  may  be  used.  Of  course,  the  ore  has  to 
contain  sufficient  sulphur  —  not  less  than  8  to  10  per  cent. 

The  reverberatory  furnaces  at  Sombrerete  roasted  from  60  to 
80  tons  of  ore  per  day,  and  the  space  available  as  a  cooling  floor 
was  inconveniently  small  for  the  regular  work,  and  a  repetition 
of  the  experiment  in  a  proper  kiln  was,  therefore,  not  practicable. 

CHLORIDIZING  ROASTING  WITH  STEAM 

If  steam  is  admitted  into  the  furnace  during  the  chloridizing 
period,  it  forms  hydrochloric  acid,  which  decomposes  the  sul- 
phides, expels  arsenic  and  antimony,  and  chloridizes  the  silver 
with  great  energy,  even  metallic  silver,  on  which  chlorine  acts 
but  slowly.  It  acts  also  on  metal  silicates  and  chloridizes  the 
silver  contained  therein,  which  otherwise  would  remain  entirely 
indifferent  to  the  action  of  the  chlorine.  The  heavy  zinc-lead  ore 
of  the  San  Francisco  del  Oro  mine,  Chihuahua,  Mexico,  when 
in  course  of  experiments  it  was  passed  through  the  Stetefeldt 
furnace,  showed  a  chlorination  of  only  15  to  16  per  cent. 
It  still  contained  8  per  cent,  sulphur,  and  in  order  to  bring  the  ore 
in  better  condition  for  the  extraction  of  the  silver  it  was  re- 
roasted.  The  chlorination,  however,  could  only  be  increased  to 
about  44.2  per  cent.  On  examination  of  the  ore  as  it  came  from 
the  Stetefeldt  furnace  it  was  found  that  by  dropping  through  the 
shaft  the  main  portion  of  the  ore  was  transformed  into  minute 
globules,  which  showed  that  the  ore  was  partially  slagged,  and 


32  HYDROMETALLURGY  OF  SILVER 

the  silver  contained  in  these  globules  resisted  the  action  of  the 
chlorine.  After  reroasting,  these  globules  felt  between  the 
fingers  just  as  sharp  and  glassy  as  before,  but  when  the  reroasting 
was  done  in  presence  of  steam  the  result  was  different.  The 
chlorination  increased  from  15  to  66.6  per  cent,  and  the  globules 
became  soft  and  could  be  powdered  between  the  fingers.  Of 
course,  a  chlorination  of  66.6  per  cent,  is  very  inferior,  but  this 
fact  does  not  interest  us  just  now.  The  ore  had  been  spoiled  in 
the  Stetefeldt  furnace,  which  made  it  impossible  to  produce  a 
satisfactory  chlorination.  The  present  interest  is  the  fact  that 
these  experiments  demonstrated  the  beneficial  effect  of  steam 
in  roasting.  Without  steam  the  chlorination  was  only  44.2  per 
cent,  while  with  steam  it  was  66.2  per  cent.  These  figures  repre- 
sent the  average  of  a  considerable  number  of  charges.  The 
globules  which  remained  unchanged  when  roasted  without  steam 
became  soft  and  assumed  the  color  of  roasted  ore. 

The  same  conditions  were  maintained  in  both  cases,  with 
regard  to  the  percentage  of  salt  to  the  temperature  applied,  etc. 
The  improved  results  can  therefore  be  credited  solely  to  the 
action  of  the  steam . 

If  an  ore  is  rich  in  lead  it  is  hardly  possible  to  avoid  the  for- 
mation of  lead  silicate  during  the  oxidizing  period,  and  the  silver 
contained  therein  will  not  be  chloridized  during  the  subsequent 
chloridizing  period,  and  consequently  will  enrich  the  residues 
and  be  lost.  Roasting  with  steam  is,  therefore,  much  to  be 
recommended  for  ores  containing  galena,  especially  if  the  galena 
is  rich  in  silver,  which  is  very  often  the  case  in  complex  ores. 

Objections  have  been  frequently  made  against  roasting  with 
steam,  based  on  the  assumption  that  it  much  increases  the  consump- 
tion of  fuel,  but  in  actual  practice  it  will  be  found  that  the  increased 
consumption  is  not  serious  at  all.  Waste  steam  from  the  engine 
can  be  used,  but  even  if  live  steam  is  applied  it  is  not  necessary 
to  use  it  in  such  volumes  as  to  cause  a  marked  drop  in  the  tem- 
perature of  the  furnace.  A  moderate  application  answers  the 
purpose.  Sometimes  the  mere  keeping  of  water  in  the  ash-pit 
has  a  decidedly  beneficial  effect.  The  steam  has  to  enter  the 
furnace  at  the  fire  end  and  under  the  flame,  so  that  it  comes 
well  in  contact  with  the  ore.  The  steam  becomes  superheated 
mostly  at  the  expense  of  the  ore  next  to  the  fire-bridge,  thus 
preventing  an  overheating  of  that  part  of  the  charge. 


METHODS  OF   ROASTING  33 

The  use  of  steam  may  also  greatly  reduce  the  loss  of  silver  by 
volatilization.  This  is  mostly  noticeable  with  ore  containing  rich 
antimonial  fahlerz  and  zinc  blende  besides  antimonial  galena. 
In  working  the  ores  of  the  Silver  King  mine  of  Arizona,  I  had 
the  opportunity  to  make  very  interesting  observations  with  regard 
to  the  effect  of  steam  in  reducing  the  loss  of  silver  by  volatilization. 

As  steam  has  not  on  all  kinds  of  ore  such  a  striking  effect  as 
in  this  case,  it  will  be  instructive  to  give  a  short  description  of 
the  Silver  King  ore.  This  was  a  complex  ore,  and  consisted  of 
the  following  silver-bearing  minerals: 

(1)  Native  silver  in  close  contact  with  fahlerz,  silver  copper 
glance,  zinc  blende,  and  in  some  instances  with  galena.     It  was 
brittle  enough,  so  that  a  large  part  of  it  was  pulverized  in  the 
battery.     It  occurred  in  the  shape  of  wire,  flakes,  solid  grains,  and 
in  large  chunks,  and  in  such  quantities  that  it  had  to  be  removed 
from  the  mortars  of  the  battery  twice  a  week  by  means  of  shovels. 
This  silver  was  of  a  bright  white  color,  0.975  fine,  and  did  not 
contain  any  gold. 

(2)  Silver  copper  glance  with  70.3  per  cent,  silver,  9.8  per  cent, 
copper,  17.4  per  cent,  sulphur. 

(3)  Antimonious  fahlerz,  containing  over  3000  oz.  of  silver 
per  ton.  This  mineral  was  the  most  important  constituent  part 
of  the  ore. 

(4)  Zinc  blende,  of  which  there  were  three  varieties: 

(a)  Zinc  blende  found  in  large  and  quite  transparent  crystals 
of  a  lustrous  green  color.     This  was  the  poorest  of  the  silver- 
bearing  minerals  of  the  ore,  but  it  was  highly  interesting  from  its 
beauty  as  a  specimen.     It  contained  only  10.2  oz.  silver  per  ton. 

(b)  Brown  zinc  blende  occurred  in  solid  masses  and  in  large 
quantities,    frequently   permeated   with   wire    silver,    and    con- 
tained 97.7  oz.  silver  per  ton. 

(c)  Black  zinc  blende  was  more  scarce,  and  contained  40.8  oz. 
silver  per  ton. 

(5)  Galena  occurred  in  two  varieties:  the  fine-grained  anti- 
monious  with  185  oz.,  and  the  coarsely  crystallized  with  only 
29  oz.  silver  per  ton. 

(6)  Peacock  copper  ore,  with  450.6  oz.  silver  per  ton. 

(7)  Copper  pyrites. 

(8)  Iron  pyrites. 

The  gangue  consisted  of  quartz,  heavyspar  and  some  por- 


34  HYDROMETALLURGY  OF  SILVER 

phyry.  The  average  value  of  the  ore  as  furnished  to  the  mill 
was  161.4  oz.  per  ton. 

This  ore  was  roasted  with  10  per  cent,  of  salt,  in  a  large  size 
revolving  furnace  of  the  Bruckner  type,  but  with  a  modification 
specially  designed  by  myself  for  this  ore.  On  account  of  the 
antimonious  fahlerz,  the  antimonious  galena  and  the  heavyspar, 
the  ore  caked  very  easily.  For  this  reason,  and  to  avoid  excessive 
loss  of  silver  by  the  antimony,  the  ore  had  to  be  roasted  at  a 
very  moderate  heat.  The  furnaces  were  16  ft.  long.  It  was 
found  that  the  roasting  could  not  be  done  properly  with  a  furnace 
of  common  construction,  with  a  fireplace  only  at  one  end  of  the 
cylinder,  as  the  ore  either  did  not  receive  enough  heat  at  the 
farther  end,  or,  if  it  did,  it  was  overheated  and  caked  at  the  end 
nearest  to  the  fire.  To  overcome  this  difficulty  the  cylinder  was 
provided  at  each  end  with  a  fireplace  and  flue  arrangement  (see 
Figs.  15,  16,  and  17).  These  two  fireplaces  were  worked  alter- 
nately. After  the  ore  was  charged,  the  furnace  was  set  in  slow 
revolving  motion,  and  fire  kept  up  in  one  of  the  fireplaces.  The 
flame  traversed  the  furnace,  and  smoke  and  gases  escaped 
through  the  flue,  in  front  of  the  opposite  fireplace.  After  a  lapse 
of  one  hour,  fire  was  made  up  in  the  other  fireplace,  the  damper 
reversed,  and  flame  and  gases  allowed  to  pass  through  the  fur- 
nace in  the  opposite  direction.  The  changing  of  the  fire  was 
kept  up  during  the  whole  time  the  charge  was  in  the  furnace, 
only  the  intervals  were  not  quite  as  frequent  as  in  the  begin- 
ning. This  system  of  double  fireplace  and  flues  proved  to  be 
of  great  advantage  in  securing  a  very  uniform  roasting;  the 
ore  from  both  ends  was  chloridized  up  to  the  same  percentage, 
while,  when  the  ore  was  roasted  in  a  furnace  with  a  fireplace  at 
one  end  only,  the  farther  end  showed  a  less  chlorination  of 
from  5  to  10  per  cent.  Besides,  it  enabled  the  operator  to 
roast  at  a  low  and  uniform  heat. 

Notwithstanding  the  capacity  of  the  furnaces  to  roast  at  a 
moderate  and  uniform  heat,  the  loss  of  silver  by  volatilization 
proved  to  be  exorbitant,  being  not  less  than  38  per  cent.,  while  at 
the  same  time  the  chlorination  was  low  on  account  of  the  large 
percentage  of  metallic  silver  in  the  ore,  which  was  but  imperfectly 
converted  into  silver  chloride  by  the  chlorine.  A  jet  of  steam 
was  then  tried,  which  was  applied  right  under  the  flame  and 
directed  toward  the  side  where  the  ore  was.  There  was  a  steam 


METHODS   OF   ROASTING  35 

jet  at  each  end,  but  only  the  one  was  operated  which  corresponded 
with  the  end  at  which  was  the  fire.  The  roasting  results  thus 
obtained  were  very  satisfactory.  An  average  of  many  charges 
showed  that  the  loss  by  volatilization  was  reduced  from  38  per 
cent,  to  2  per  cent,  while  the  average  chlorination  of  67  furnace 
charges  proved  to  be  94.4  per  cent,  and  in  some  cases  as  high  as 
96.8  per  cent. 

This  roasting  example  illustrates  that  with  certain  ores  the 
application  of  steam  is  of  vital  importance.  The  ores  of  the 
Silver  King  mines  could  not  have  been  worked  by  a  hydro- 
metallurgical  method  if  the  steam  had  not  so  greatly  reduced 
the  loss  of  silver,  and  increased  the  percentage  of  chlorination. 
There  are  some  ores  which  do  not  need  steam,  but  in  most 
cases  a  larger  or  smaller  jet  of  steam,  according  to  the  nature  of 
the  ore,  does  beneficially  assist  the  chemical  reactions. 

CHLORIDIZING  ROASTING  OF  SILVER  ORES 
CONTAINING  GOLD 

There  are  two  combinations  of  gold  and  chlorine:  the  aurous 
and  the  auric  chloride.  The  latter  is  soluble  in  water  and  is 
formed  when  finely  divided  gold  is  brought  in  contact  with 
chlorine  gas  at  a  common  or  moderately  warm  temperature. 
At  a  temperature  of  230  deg.  C.  it  changes  into  aurous  chloride, 
which,  however,  on  further  heating,  decomposes  into  metallic 
gold  and  chlorine.  Owing  to  this  property  of  the  chlorino  com- 
pounds of  the  gold,  neither  of  them  will  be  formed  in  the  furnace 
during  chloridizing  roasting.  The  temperature  in  the  furnace 
is  too  high  for  them  to  exist,  and  the  gold  on  discharge  of  the 
furnace  will  be  found  in  the  metallic  state.  The  aurous  chloride, 
which  is  not  soluble  in  water  but  is  soluble  in  a  solution  of  sodium 
hyposulphite,  resists  a  much  higher  temperature  than  does  the 
auric  chloride,  and  it  will  form  at  a  temperature  much  higher 
than  the  decomposing  point  of  the  auric,  which  temperature, 
however,  has  to  be  kept  below  red  heat. 

Based  on  this  property  of  the  gold  chlorides  I  adopted  a  modus 
operandi  by  which  I  was  able  to  extract  75  to  80  and  even  90 
per  cent,  of  the  gold  contained  in  the  silver  ore  simultaneously 
with  the  silver  by  sodium  hyposulphite. 

If  the  ore  leaving  the  furnace  is  not  allowed  to  cool  quickly, 


36  HYDROMETALLURGY  OF  SILVER 

but,  on  the  contrary,  is  made  to  cool  slowly  by  dumping  it  to  a 
large  pile  and  leaving  it  undisturbed  until  it  is  cool,  which  takes 
several  days,  it  will  be  observed  that  the  generation  of  chlorine 
still  continues  for  a  considerable  time.  The  cooling  of  the  heap 
begins  from  the  outside  and  progresses  toward  the  inside,  and 
the  chlorine,  which  is  generated  at  the  inside,  in  escaping  will 
meet  a  layer  of  ore  sufficiently  cooled  for  combination  with  the 
gold  contained  therein.  It  will  form  the  aurous  chloride,  because 
the  temperature  is  still  too  high  for  the  auric  chloride  to  exist. 
The  formation  of  aurous  chloride  will  progress  toward  the  inside 
in  proportion  to  the  cooling  of  the  pile.  The  cooling  does  not 
need  to  be  continued  beyond  the  time  when  a  sample  taken  from 
the  inside  does  not  emit  any  chlorine. 

If  no  precaution  is  taken  to  cool  the  ore  slowly,  only  a  small 
percentage  of  the  gold  will  be  converted  into  aurous  chloride, 
and  the  gold  extraction,  therefore,  will  be  very  small.  The 
beneficial  effect  of  slow  cooling  on  the  chlorination  of  the  gold 
contained  in  auriferous  silver  ore  can  also  be  observed  in  ex- 
perimenting on  a  small  scale,  which  will  be  illustrated  by  some 
results  which  I  recently  obtained  in  conducting  some  laboratory 
investigations  respecting  the  ore  of  the  Lucky  Tiger  mine,  Sonora, 
Mexico.  An  analysis  of  the  sample  showed  the  ore  to  consist  of: 

Iron 2.92  per  cent. 

Zinc 3.36 

Lead 1.15 

Copper trace. 

Antimony trace. 

Sulphur 2.54  per  cent. 

Silica 89.54 

Silver 108.16  ounces  per  ton. 

Gold 0.36 

Two  lots  of  100  grams  each  were  roasted  with  salt  on 
roasting  dishes  in  the  muffle  at  a  dark-red  heat.  The  amount  of 
salt  as  well  as  the  temperature  and  roasting  time  were  for  both 
lots  exactly  the  same.  When  roasting  was  completed  one  lot 
was  withdrawn  and  allowed  to  cool  at  a  place  away  from  the 
muffle,  while  the  other  was  placed  in  a  hot  roasting-dish  and 
covered  with  another  hot  roasting-dish,  then  removed  from  the 
muffle,  but  placed  right  in  front  of  it.  Thus  the  one  lot  was 
allowed  to  cool  quickly,  while  the  other  was  made  to  cool  slowly. 
The  quickly  cooled  ore  showed  a  gold  chlorination  of  20.8  per 
cent,  that  is,  20.8  per  cent,  of  it  could  be  extracted  with  sodium 


METHODS   OF   ROASTING  37 

hyposulphite,  while  the  slowly  cooled  ore  showed  a  gold  chlori- 
nation  of  74.7  per  cent.  The  result  of  this  experiment  clearly 
demonstrates  that  the  chlorination  of  the  gold  takes  place  out- 
side the  furnace  and  is  caused  by  slow  and  gradual  cooling.  The 
conditions  in  this  experiment  were  not  as  favorable  as  they 
would  have  been  on  a  large  scale,  because  the  generation  of 
chlorine  in  so  small  a  lot  as  100  grams  ceases  soon,  while  in  a 
large  pile  it  continues  for  many  hours. 

While  this  mode  of  roasting  gives  very  satisfactory  results 
with  silver  ores  containing  1  oz.  of  gold  per  ton  or  less,  it  will 
not  be  quite  satisfactory  for  ores  richer  in  gold,  in  which  case 
a  cyanide  solution  should  be  applied  after  the  extraction  of  the 
silver.  In  the  above  experiment  the  residues  of  the  quickly 
cooled  ore  were  treated  with  a  cyanide  solution,  by  which  the 
gold  extraction  was  raised  from  20.8  per  cent  to  86.07  per  cent. 

CHLORIDIZING  ROASTING  FOR  AMALGAMATION 

We  have  seen  that  the  aim  in  chloridizing  roasting  for  the 
extraction  of  the  silver  by  lixiviation  with  sodium  hyposulphite 
is  to  convert  as  much  as  possible  of  the  silver  into  silver  chloride 
and  at  the  same  time  to  expel  as  little  as  possible  of  the  volatile 
base-metal  chlorides,  in  order  to  reduce  the  loss  of  silver  by  vola- 
tilization to  the  minimum.  This  modification  in  chloridizing  roast- 
ing was  an  important  step  forward  in  the  hydrometallurgy  of 
silver  because  thereby  its  weakest  and  most  objectionable  feature, 
the  volatilization  of  the  silver,  was  reduced  to  the  minimum. 

Chloridizing  roasting  was  first  devised,  studied,  and  executed 
to  meet  the  requirements  of  barrel  and  later  of  pan  amalgamation. 
For  this  process  not  only  as  much  as  possible  of  the  silver  has 
to  be  converted  into  silver  chloride,  but  it  is  of  the  greatest  im- 
portance to  expel  or  to  decompose  the  base-metal  chlorides,  too, 
because  if  this  is  not  done  these  chlorides  will  amalgamate  with 
the  mercury  as  well  as  the  silver  and  produce  an  impracticably 
large  amount  of  amalgam  and  a  very  low-grade  bullion.  Besides 
this,  the  mercury  containing  much  of  such  amalgam  has  no  ten- 
dency to  unite  when  the  pulp  is  diluted,  but  remains  in  minute 
globules,  which  partly  float  on  the  surface  of  the  pulp  as  a  dark 
scum  and  partly  are  carried  off  with  the  residue,  thus  causing  a  very 
large  loss  of  mercury  and,  of  course,  of  the  silver  it  contains. 


38  HYDROMETALLURGY  OF  SILVER 

The  mercury  loses  much  of  its  decomposing  energy  on  silver 
chloride;  it  becomes  "foul/7  which  results  in  very  rich  residues. 
In  short,  it  is  absolutely  necessary  in  working  complex  silver  ores 
by  amalgamation  to  expel  and  decompose  the  base-metal  chlorides. 
This  is  done  by  heat. 

The  salt  is  added  either  together  with  the  ore  or  after  the 
oxidation  of  the  sulphides  has  pretty  well  advanced.  During 
the  oxidizing  period  the  temperature  is  kept  moderate,  partly  to 
prevent  caking  but  mostly  to  form  as  much  metal  sulphates  as 
possible,  because  these  sulphates  will  act  on  the  salt  and  change 
into  chlorides  and  also  produce  chlorine,  and,  in  presence  of  steam, 
hydrochloric  acid.  If  the  ore  is  rich  in  sulphurets  the  heat 
created  by  their  oxidation  can  so  increase  as  to  make  it  necessary 
to  let  the  fire  go  out  entirely,  but  close  attention  has  to  be  paid 
to  start  the  fire  again  as  soon  as  a  pronounced  decrease  of  the 
temperature  can  be  observed.  If  an  ore  charge  is  allowed  to  cool 
too  much  it  takes  considerable  time  and  fuel  to  bring  the  temper- 
ature up  again  to  the  desired  degree.  During  the  oxidizing 
period  the  ore  ought  to  be  frequently  stirred  —  continually,  if 
possible.  The  beginning  of  the  chloridizing  period  is  indicated 
by  the  swelling  of  the  ore;  it  becomes  "woolly"  and  emits  heavy 
white  fumes.  The  heat  is  then  gradually  increased  until  toward 
the  end  the  charge  assumes  a  bright  cherry  red.  During  roasting 
the  charge  is  several  times  turned,  that  is,  the  half  toward  the 
fire-bridge  is  removed  toward  the  flue,  and  that  from  the  flue 
side  is  brought  toward  the  fire-bridge.  This  is  done  in  the 
following  way:  The  whole  charge  is  raked  together  into  a  long 
heap  extending  from  the  bridge  to  the  flue  end.  This  ridge  of 
ore  is  made  nearer  to  the  working  doors,  to  lay  bare  as  much  of 
the  hearth  as  possible.  Now,  by  means  of  a  shovel  the  ore  from 
the  fire-bridge  is  moved  and  dumped  back  of  the  ridge,  com- 
mencing close  at  the  flue  and  proceeding  toward  the  fire-bridge 
until  the  middle  is  reached,  then  the  other  part  of  the  ridge  is 
moved  toward  the  bridge.  Then  the  whole  charge  is  spread 
again. 

During  roasting,  samples  are  taken  and  tested  with  regard  to 
the  behavior  of  the  ore  toward  mercury.  If  it  is  found  that  by 
adding  some  water  and  some  mercury  to  the  ore  the  bright  sur- 
face of  the  latter  becomes  immediately  covered  with  a  black 
skin,  it  is  an  indication  that  the  roasted  ore  still  contains  an 


METHODS  OF  ROASTING  39 

injurious  amount  of  base-metal  chlorides,  and  roasting  has  to  be 
continued  at  an  increased  temperature  in  order  to  decompose 
them. 

The  practice  of  decomposing  the  base-metal  chlorides  by  heat 
and  increased  roasting  time  is  naturally  connected  with  much  loss 
of  silver  by  volatilization.  The  late  G.  Kiistel  proposed  a  much 
more  rational  means  than  heat  to  destroy  the  base-metal  chloride. 
In  his  treatise,  "Roasting  of  Gold  and  Silver  Ores,"  he  says: 

"It  is  very  difficult  to  get  rid  of  all  the  base  chlorides;  they 
are  formed  under  the  action  of  chlorine  and  hydrochloric  acid. 
The  most  of  the  metal  chlorides  are  volatile,  and  a  part  is  carried 
off  through  the  chimney.  Another  part  of  the  chlorides  gives 
off  some  of  its  chlorine,  whereby  sulphates,  undecomposed  sul- 
phurets,  antimonates,  and  arsenates  are  chloridized.  Chlorides 
which  are  disposed  to  transfer  chlorine  to  other  metals  in  com- 
bination with  sulphur  or  arsenic  are:  the  proto  chloride  of  iron 
and  of  copper,  the  chlorides  of  zinc,  lead,  and  cobalt.  When  in 
this  way  the  most  of  the  metals  are  chloridized,  the  base  metals, 
principally  iron  and  copper,  are  losing  their  chlorine  gradually, 
being  first  converted  into  sub-chlorides,  and  then  into  oxides. 
The  roasting  for  this  purpose  must  continue  with  increased  heat, 
even  when  the  chlorination  of  the  silver  is  finished.  At  an 
increased  heat,  the  base  metal  chlorides  lose  their  chlorine,  while 
the  chloride  of  silver  remains  undecomposed,  unless  a  very  high 
temperature  should  be  applied.  This  process  requires  a  long 
time,  consequently  also  more  fuel.  The  decomposition  of  these 
chlorides  is  greatly  assisted  by  the  use  of  5  to  6  per  cent,  of  carbo- 
nate of  lime  in  a  pulverized  condition.  Lime  does  not  attack  the 
chloride  of  silver,  but  it  is  not  advisable  to  take  too  much  of  it, 
as  it  would  interfere  to  some  degree  with  the  amalgamation. 
The  pulverized  lime-rock  must  be  charged  toward  the  end  of 
the  roasting.  First,  two  per  cent,  is  introduced  by  means  of  a 
scoop,  the  whole  well  mixed,  and  then  examined  either  with 
sulphide  of  sodium  or  in  the  following  way: 

"A  small  portion  of  the  roasted  ore  is  taken  in  a  porcelain 
cup  or  glass,  and  mixed  with  some  water  by  means  of  a  piece  of 
iron  with  a  clean  metallic  surface.  If  the  iron  appears  coated 
red  with  copper,  some  more  lime  must  be  added.  In  place  of 
iron — especially  if  no  copper,  but  some  other  base  metal  is 
present  —  some  quicksilver  is  mixed  with  the  sample.  In  the 


40  HYDROMETALLURGY  OF  SILVER 

presence  of  base-metal  chlorides,  the  quicksilver  is  coated  imme- 
diately with  a  black  skin. 

"When  endeavoring  to  expel  the  base  metals  by  heat,  the 
loss  of  silver,  in  presence  of  much  antimony,  lead,  and  copper, 
should  be  investigated  very  carefully.  Under  certain  circum- 
stances it  is  not  uncommon  to  find  a  loss  of  even  50  per  cent,  of 
the  silver,  if  the  chloridizing  roasting  is  carried  on  at  a  high  heat 
for  a  great  length  of  time.  The  loss  increases  with  the  duration 
of  roasting  and  with  the  degree  of  temperature.  When  such  ore 
is  under  treatment,  it  is  necessary  to  take  samples  during  the 
roasting,  and  to  examine  the  same  for  the  amount  of  chloride  of 
silver,  and  also  for  its  loss,  and  to  stop  roasting  when  the  highest 
percentage  of  chloride  of  silver  is  obtained  without  reference  to 
the  condition  of  base  metals." 

Mr.  Kiistel  proposed  a  still  more  rational  method  for  remov- 
ing from  the  roasted  ore  the  base-metal  chlorides  which  are  so 
obnoxious  to  amalgamation.  Instead  of  destroying  them  by 
heat  and  extended  roasting  time,  he  removed  all  soluble  chlorides 
and  sulphate  of  zinc  by  leaching  the  ore  with  hot  water  previous 
to  amalgamation.  The  ores  at  Flint,  Idaho,  turned  out  such 
base  amalgam  that  further  working  proved  to  be  ruinous;  but 
after  Mr.  Kiistel  applied  his  method  the  ore  became  most  suit- 
able for  amalgamation,  and  very  satisfactory  results  were 
obtained. 

It  is  apparent  that  this  method  is  quite  an  advance  in 
amalgamation;  not  only  is  the  amalgamating  energy  between 
silver  and  mercury  much  increased,  thus  resulting  in  a  better 
extraction  of  the  silver,  but  the  volatilization  of  the  silver  in 
the  furnace,  and  the  loss  of  mercury  and  silver  in  the  process  of 
amalgamation,  are  much  reduced. 

An  important  point  in  this  process  should  not  be  overlooked, 
namely,  that  chloride  of  silver,  while  not  soluble  in  water,  is 
soluble  in  a  concentrated  solution  of  metal  chlorides.  The  dis- 
solving energy  of  such  a  solution  on  silver  chloride  increases  with 
its  concentration  and  its  temperature.  Hot  water  has  to  be 
used  in  order  to  remove  the  lead  chloride,  and  therefore  the 
first  part  of  the  outflowing  solution  will  be  as  a  rule  rather  con- 
centrated, and  at  the  same  time,  being  warm,  will  dissolve  quite 
a  noticeable  amount  of  silver  chloride.  This,  however,  can  be 
regained  by  collecting  the  solution  in  large  tanks  and  by  diluting 


METHODS  OF  ROASTING  41 

the  same.  If  sufficient  water  is  added,  all  the  silver  chloride  will 
precipitate,  and  if  enough  time  can  be  given  will  settle  on  the 
bottom  together  with  lead  chloride.  If  there  is  copper  in  the 
ore  the  solution  should  be  made  to  flow  through  tanks  filled  with 
scrap  iron,  by  which  the  copper  and  silver  are  saved. 

Lead  chloride  amalgamates;  lead  sulphate  does  not;  and  in 
roasting  for  amalgamation  it  is  therefore  of  importance  to  con- 
vert as  much  of  the  lead  as  possible  into  sulphate  and  as  little  of 
it  as  possible  into  chloride. 

Lead  sulphate,  if  once  formed,  remains  indifferent  and  un- 
changed during  the  balance  of  the  process.  Mr.  Kiistel,  in  roast- 
ing the  plumbiferous  silver  ores  for  amalgamation  at  Plomosas, 
Mexico,  made  the  observation  that  if  the  ore  was  roasted  in  a 
reverberatory  with  a  roof  of  only  20  or  less  inches  above  the  roast- 
ing floor  the  bullion  obtained  contained  500  to  600  parts  of  lead 
in  1000,  while  if  the  same  ore  was  roasted  in  a  reverberatory  with 
high  roof,  27  or  30  inches,  the  bullion  obtained  was  almost  free 
of  lead.  In  the  first  instance  much  lead  chloride  and  oxichloride 
was  formed,  which  amalgamated,  while  in  the  second  instance 
nearly  all  the  lead  was  transformed  into  lead  sulphate.  The 
reason  of  the  different  results  is  apparent.  In  the  furnace  with 
the  higher  roof  there  was  sufficient  space  left  between  the  flame, 
which  travels  next  to  the  roof,  and  the  surface  of  the  ore  to  per- 
mit a  free  access  of  air,  and  in  presence  of  ample  live  air  more 
sulphuric  acid  is  formed,  which  acts  on  lead  oxide,  oxichloride, 
and  chloride,  converting  them  into  sulphate.  If  the  supply  of  air 
is  limited  by  the  low  arch,  insufficient  sulphuric  acid  is  formed 
to  convert  all  the  lead  into  sulphate.  In  support  of  this  theory 
speaks  the  fact  that  C.  A.  Stetefeldt,  by  roasting  the  ore  from 
Ontario,  Utah,  in  a  Stetefeldt  furnace,  found  all  the  lead  contained 
in  the  ore  to  be  in  the  state  of  sulphate. 

With  regard  to  the  extraction  of  the  silver  by  sodium  hypo- 
sulphite it  is  immaterial  whether  the  lead  in  the  roasted  ore  is  in 
the  state  of  sulphate  or  chloride.  If  cold  water  is  used  for  re- 
moving the  base  metal  chlorides,  which  is  generally  the  case, 
but  a  small  portion  of  the  lead  chloride  is  removed,  and  the  same 
is  still  contained  in  the  ore  when  the  latter  is  subjected  to  leach- 
ing with  sodium  hyposulphite,  in  which  solution  it  is  soluble,  as 
well  as  the  sulphate,  which  is  not  soluble  in  water. 


VI 

CONSUMPTION  OF  FUEL 

THE  fuel  mostly  used  in  roasting  is  wood,  not  so  much  so 
because  it  is  the  fuel  with  which  it  is  the  easiest  to  regulate  the 
temperature,  but  because  the  mines  and  works  are  usually  situated 
in  more  or  less  remote  mining  districts,  where  wood  is  easier  and 
cheaper  to  be  procured  than  other  fuel.  Bituminous  coal  and 
gas  from  gas  producers  answer  the  purpose  perfectly  well. 

The  quantity  of  fuel  required  depends  mostly  on  the  nature 
of  the  ore,  but  also  on  the  construction  of  the  furnace.  Ores 
rich  in  sulphur,  especially  if  a  large  part  of  it  is  combined  with 
iron,  as  in  iron  pyrites,  require  the  least  amount,  because  by 
the  combustion  of  the  sulphur  much  heat  is  developed  in  the  ore 
itself,  so  that  the  process  needs  only  to  be  slightly  assisted  by 
fire.  In  fact,  ores  with  20  or  22  per  cent,  sulphur,  if  roasted  in  a 
Bruckner  furnace,  need  only  to  be  heated  until  it  is  ignited.  The 
balance  of  the  heat  required  for  the  oxidation  and  chlorination 
is  then  furnished  by  the  ore,  the  same  as  in  chloridizing  self-roast- 
ing, by  which  method  10  or  more  tons  of  ore  can  be  roasted  with 
one  cord  of  wood.  A  highly  sulphureted  ore  which  does  not  con- 
tain too  much  zinc  blende  or  galena  but  contains  considerable 
pyrite  could  be  roasted  chloridizingly  without  the  use  of  any  fuel 
except  what  is  needed  to  start  the  furnace.  We  have  seen  above, 
in  heap-roasting,  that  only  a  partial  oxidation  of  the  ore  in  the 
furnace  is  required;  and  if  for  this  purpose  a  continually  dis- 
charging furnace  of  suitable  construction  is  used,  like  those  of 
the  McDougal  type  as  employed  in  sulphuric  acid  factories,  or  of 
the  Howell  type  (if  the  latter  is  lined  with  bricks  the  whole 
length  and  only  a  very  slight  inclination  given,  and  provided 
with  a  ring-flange  at  both  ends  so  as  to  retain  a  somewhat  larger 
amount  of  ore  in  the  furnace  in  order  to  produce  more  heat  and 
to  preserve  it  better)  a  continuous  chloridizing  roasting  could  be 
successfully  effected  without  the  use  of  fuel. 

These  continually  discharging  furnaces  will  have  to  discharge 

42 


CONSUMPTION  OF   FUEL  43 

into  closed  pits,  which  in  starting  will  have  to  be  first  well  heated 
like  the  furnaces.  From  these  pits  or  vaults  the  ore  is  then 
moved  and  charged  into  the  roasting  bins.  The  vaults  should  be 
sufficiently  large  to  prevent  cooling. 

Ores,  very  poor  in  sulphur,  containing  only  about  2  per  cent., 
require  much  fuel,  and  if  the  same  is  expensive  should  not  be 
roasted  in  a  reverberatory  furnace,  because  the  conditions  pre- 
vailing in  a  reverberatory  furnace  as  regards  utilization  of  heat 
are  very  unfavorable.  Heat  .penetrates  but  very  slowly  into  any 
pulverized  material,  especially  if  the  same  is  left  undisturbed. 
If  a  fresh  charge  is  placed  on  the  hearth  of  a  reverberatory  and 
spread  over  it,  it  will  cool  the  bottom,  and  if  it  is  not  a  muffle 
furnace,  the  supply  of  heat  from  that  source  will  soon  end,  and 
the  heating  will  be  effected  only  on  the  surface  by  the  radiating 
heat  of  the  flame  which  travels  next  to  the  roof  of  the  furnace; 
but  heat,  as  stated,  penetrates  very  slowly  into  pulverized  ma- 
terial, and  therefore  only  a  small  percentage  of  the  heat  produced 
by  the  fuel  will  be  effective.  To  better  utilize  the  heat  it  is 
necessary  that  continually  new  particles  of  the  ore  be  brought  to 
the  surface  to  be  exposed  to  the  radiating  heat.  This,  however, 
for  obvious  reasons,  can  be  done  but  very  imperfectly  in  a  hand- 
stirred  furnace.  As  soon  as  the  charge  becomes  hotter  than  the 
bottom  of  the  furnace  the  bottom  will  draw  heat  from  the  charge, 
thus  exerting  a  cooling  action  on  the  latter.  To  make  the  furnace 
long  helps,  but  not  very  much;  and  if  the  roasting  of  such  dry 
ore  has  to  be  done  in  a  reverberatory  furnace  it  is  more  rational 
to  make  the  hearth  short  and  to  build  the  furnace  in  two  stories, 
in  which  case  the  roof  of  the  lower  hearth  will  heat  the  bottom 
of  the  upper  one. 

The  most  effective  and  at  the  same  time  the  most  fuel-saving 
roasting  furnace  for  this  class  of  ores  is  undoubtedly  the  Stetefeldt 
furnace,  in  which  the  flame  is  made  to  ascend  through  a  high 
shaft,  while  the  ore  in  a  fine  shower  falls  down  the  shaft  and 
against  the  flame.  There  is  no  other  roasting  furnace  in  which 
the  heat  is  utilized  to  such  advantage  as  in  the  Stetefeldt ;  ten  to 
twelve  tons  of  such  dry  ore  can  be  roasted  with  one  cord  of  wood  — 
a  result  which  cannot  be  obtained  in  a  reverberatory  furnace. 
This  peculiarity  of  the  Stetefeldt  furnace,  which  is  so  advantageous 
for  dry  ore,  makes  it  also  unsuitable  for  ores  heavily  charged  with 
sulphurets. 


44  HYDROMETALLURGY  OF  SILVER 

If  circumstances  do  not  allow  the  erection  of  a  Stetefeldt 
furnace,  the  next  best  fuel-economizing  roasting  furnace  for  such 
ores  is  the  Howell -White. 

If  the  ore  permits  a  closer  sorting  without  causing  too  much 
waste  of  silver,  it  will  be  well  to  do  so  in  order  to  raise  the  sulphur 
contents  of  the  ore,  because  this  will  not  only  reduce  the  consump- 
tion of  fuel,  but  also  improve  the  extraction  and  shorten  the 
roasting  time.  In  case  the  ore  is  not  very  rich,  and  concentrates 
well,  it  may  be  advisable  to  concentrate  part  of  it  and  to  add  the 
concentrates  to  the  balance  of  the  ore,  or  to  sort  close  and  to 
concentrate  the  second-class  ore,  just  as,  according  to  circum- 
stances, is  found  to  be  the  most  rational. 


VII 


REVERBERATORY  FURNACES  WORKED  BY  HAND 

THE  reverberatory  furnace  is  a  horizontal  hearth  furnace 
provided  with  a  fireplace  and  grate  at  one  end  and  a  flue  at  the 
opposite  end,  and  with  working  doors  on  one  or  on  both  of  the 
two  long  sides.  The  hearth  is  separated  from  the  fireplace  by 
the  fire-bridge.  It  is  the  oldest  and  the  most  primitive  type  of 
roasting  furnace,  but  notwithstanding  this  it  is  the  furnace 
which  can  be  applied  to  any  kind  of  ore,  except  those  the  nature 
of  which  prohibits  chloridizing  roasting  altogether,  like  ores  con- 
taining too  much  lead.  Its  construction  gives  the  operator  full 
control  over  the  process.  It  offers  facilities  to  maintain  any  con- 
dition the  nature  of  an  ore  requires;  and  it  is,  when  substantially 
built,  very  durable,  requiring  little  repair,  and  makes  but  little 
flue-dust.  The  reverberatory  is,  in  fact,  the  ideal  furnace  for 
chloridizing  roasting,  and  would  be  exclusively  used  for  this 
purpose  if  it  were  not  for  the  fact  that  it  has  to  be  operated  by 
hand,  which  makes  the  cost  of  roasting  too  high  in  localities 
where  labor  is  costly.  This  was  the  cause  which  gave  the  impulse 
in  the  silver  districts  of  the  Pacific  coast  of  the  United  States  to 
the  invention  and  construction  of  quite  a  variety  of  mechanical 
furnaces,  all  of  which  are  labor-saving,  and  if  applied  to  the 
proper  ore  do  excellent  service.  They  cannot  be  used,  however, 
for  so  many  kinds  of  ore  as  the  reverberatory.  Each  of  them 
has  its  peculiarities,  with  which  the  character  of  the  ore  has  to 
comply,  and  it  is  therefore  of  the  greatest  importance,  if  mechani- 
cal furnaces  are  to  be  erected,  that  a  thorough  study  of  the 
nature  of  the  ore  should  precede  the  selection  of  the  furnace. 

The  Single-Hearth  Reverberatory.  —  This  is  the  oldest  style  of 
a  roasting  furnace.  Figs.  1  and  2  represent  the  vertical  and  the 
horizontal  section  respectively:  a,  hearth;  s,  roof;  h,  fireplace; 
it  fire-bridge;  e  and  e',  flue;  b,  bottom  discharge  opening;  d,  vault 
for  placing  the  wheelbarrow  to  receive  the  roasted  ore;  p,  charge 

45 


46 


HYDROMETALLURGY  OF  SILVER 


hopper  in  the  roof  of  the  furnace.  (This  hopper  is  provided  with 
a  slide  which,  when  drawn,  permits  the  charge  to  drop  into  the 
furnace.  It  has  to  be  large  enough  to  hold  a  full  charge,  and 
ought  to  hang  on  a  proper  framework,  so  that  its  weight  does  not 
rest  on  the  roof  of  the  furnace) ;  w,  buck  stays  and  anchors. 


i    a    a    4? 


FIGS.  1  and  2.  —  SINGLE-HEARTH   REVERBERATORY  FURNACE. 

The  single-hearth  furnaces  are  not  in  use  any  more  for  roasting 
ore  on  a  large  scale.  They  are  too  wasteful  with  regard  to  fuel; 
the  heat  of  the  gases  is  not  utilized.  We  find  them,  however, 
very  useful  for  experimental  work  and  for  burning  the  precipitate 
in  lixiviation  works. 

The  Two-Story  Single-Hearth  Furnace.  —  This  is  a  consider- 
able improvement  on  the  single-hearth  furnace.  It  is  shown  in 


REVERBERATORY   FURNACES  WORKED  BY  HAND 


47 


Fig.  3,  which  represents  in  a  longitudinal  section  the  general 
arrangement:  a,  a  are  the  lower  and  upper  hearth ;r, lower  fireplace; 
6,  flue  connecting  the  lower  with  the  upper  hearth;  &',  flue  in  the 
arch  of  the  upper  hearth,  whence  the  gases  are  led  to  the  dust- 
chambers;  /,  working  doors  of  the  upper  hearth  (the  working 
doors  of  the  lower  hearth  are  on  the  opposite  side);  r',  an  aux- 
iliary fireplace  for  the  upper  hearth,  which  is  smaller  and  is 
used  only  when  a  fresh  charge  enters  the  furnace,  to  assist  in 
heating  and  to  ignite  the  same  more  quickly. 


FIG.  3.  —  TWO-STORY,  SINGLE-HEARTH  REVERBERATORY 
FURNACE. 

As  soon  as  the  charge  of  the  lower  hearth  is  finished  and  re- 
moved the  upper  charge  is  dumped  down  through  the  flue  6.  To 
facilitate  this  operation,  and  to  permit  the  use  of  a  hoe  instead  of 
the  slower- working  shovel,  there  is  at  the  end  side  of  the  furnace 
the  door  /,  through  which  is  inserted  the  hoe  with  which  the  charge 
is  pushed  to  drop  through  b. 

The  Long  Reverberatory  Furnace.  —  A  further  step  in  the  develop- 
ment of  the  reverberatory  furnace  was  the  construction  of  long 
furnaces  with  three  to  five  hearths  on  the  same  level,  or  in  flat 
steps  of  three  to  five  inches  rise.  The  length  depends  on  the  nature 
of  the  ore.  For  highly  sulphureted  ore,  especially  if  it  contains 
much  iron  pyrites,  the  length  may  be  extended  to  50  ft.  without 
the  aid  of  an  auxiliary  fireplace.  As  a  rule  each  hearth  is  made 
10  ft.  long  and  10  ft.  wide.  The  arch,  which  can  be  made  rather 
high  at  the  fire  end,  ought  to  slope  down  toward  the  flue  end  to 


48  HYDROMETALLURGY  OF  SILVER 

throw  the  heat  of  the  moving  gases  more  toward  the  bottom  at 
the  part  of  the  furnace  remote  from  the  fire.  If  the  arch  is  made 
straight,  each  succeeding  hearth  should  be  a  few  inches  above  the 
preceding,  by  which  the  same  object  is  attained.  These  steps 
serve  at  the  same  time  as  a  mark  for  each  hearth,  and  assist  in 
preventing  the  charges  from  getting  mixed. 

These  long  furnaces  are  either  built  singly  or  in  pairs  back 
to  back,  as  shown  by  Figs.  4,  5  and  6.  Single  furnaces  are  to 
be  preferred,  as  they  offer  the  opportunity  of  providing  work- 
ing doors  on  both  sides,  which  not  only  facilitates  the  working 
of  the  charge  but  also  permits  a  free  access  of  air  from  both  sides, 
which  is  of  great  advantage.  The  construction  of  these  single 
furnaces,  however,  is  much  more  costly,  and  at  the  same  time 
requires  a  good  deal  more  space  and  consequently  much  larger 
buildings.  A  reverberatory  furnace  with  working  doors  on  one 
side  only  has  two  dead  places,  that  is,  places  which  are  not  reached 
by  live  air,  for  which  reason  the  process  of  roasting  on  these 
places  is  not  only  retarded,  but  also  such  places  become,  as  a 
rule,  overheated  and  often  cause  there  caking  of  the  ore.  These 
places  are:  the  part  of  the  hearth  next  to  the  fire-bridge  extend- 
ing almost  to  the  first  working  door,  and  the  part  along  the  back 
wall  of  the  furnace  extending  from  the  fire-bridge  to  the  flue,  so 
that,  if  no  provision  is  made  for  air  to  enter  at  these  points,  which 
we  very  often  find  to  be  the  case,  we  have  to  consider  the  furnace 
to  be  of  faulty  construction.  The  oxygen  of  the  air  is  the  very 
life  of  the  process,  and  if  the  same  is  withheld,  or  its  access  ob- 
structed, the  bad  effect  will  invariably  manifest  itself  by  an  in- 
ferior roasting  result  and  a  higher  loss  of  silver  by  volatilization. 
These  dead  spaces  are  avoided  by  constructing  air  channels 
leading  from  the  front  under  the  hearth  and  entering  the  furnace 
through  openings  in  the  back  wall  as  shown  by  b  and  b' ',  Figs.  4 
and  6.  The  fire-bridge  is  also  provided  with  an  air  channel,  /, 
which  is  in  communication  with  three  openings,  /',  through  which 
the  air  enters  the  furnace.  The  effect  of  this  additional  air  supply 
is  very  noticeable.  The  temperature  through  the  whole  width 
of  each  hearth  is  quite  uniform  and  no  overheating  next  to  the 
fire-bridge  takes  place.  At  the  same  time  the  fire-bridge  is  much 
protected  by  the  cooling  effect  of  the  air.  If  steam  is  to  be  used 
in  roasting,  the  channel  /  serves  for  this  purpose.  The  steam 
pipe,  which  is  provided  with  three  holes  while  the  end  is  closed 


REVERBERATORY   FURNACES   WORKED   BY   HAND 


49 


—44 


50 


HYDROMETALLURGY  OF  SILVER 


by  a  cap,  is  inserted  through  /.  The  holes  in  the  pipe  are  so 
divided  that  each  one  is  located  right  under  each  upraise  of  the 
channel. 

The  long  furnace,  as  represented  by  Figs.  4,  5  and  6,  was 
designed  by  me  for  chloridizing  the  calcareous  and  arsenical  ores 
of  the  Anglo-Mexican  Mining  Company  at  Yedras,  Sinaloa, 
Mexico.  This  ore  was  of  a  very  peculiar  character,  and  apt  to 
sustain  an  unusually  heavy  loss  of  silver  if  certain  conditions  in 
the  treatment  were  not  scrupulously  maintained.  The  dimen- 
sions of  different  parts  of  this  furnace  were  designed  to  conform 
with  the  peculiarities  of  this  ore,  but  the  general  arrangement 
was  not  altered;  and  the  diagrams  will  well  serve  to  illustrate  the 
construction  of  the  long  reverberatory  furnace. 


Section  J.K. 

FIG.  6.  —  LONG  REVERBERATORY 
FURNACE. 

It  is  always  advisable  to  make  the  fireplace  (see  K,  Fig.  5) 
sufficiently  wide,  because  it  gives  the  means  to  regulate  the  tem- 
perature, and  if  too  narrow  may  make  it  impossible  to  supply 
sufficient  heat.  The  heat  of  the  first  and  second  hearths  is  mostly 
supplied  by  the  combustion  of  the  sulphides  on  the  preceding 
hearths.  Wood  requires  a  wider  fireplace  than  coal;  2  ft.  6  in. 
will  be  sufficient  for  wood  fire.  The  sides  and  roof  of  the  fire- 
box should  be  lined  with  fire-bricks  in  order  to  resist  better  the 
heat  and  the  wear  caused  by  wood  and  tools.  The  brickwork 
encasing  the  fire-box  should  be  substantial  and  well  braced. 
The  depth  should  not  be  made  much  longer  than  the  length  of 
the  wood,  which  is  usually  cut  4  ft.,  so  that  a  depth  of  5  ft.  is 
sufficient.  It  is  well  to  keep  water  in  the  ash  pit,  not  only  to 
lengthen  the  life  of  the  grate  bars  but  also  to  make  a  limited 
amount  of  steam,  which  benefits  the  roasting  very  much. 

The  top  of  the  fire-bridge  (see  m,  Fig.  5)  should  not  be 
too  high  above  the  grate,  in  order  that  the  furnace  may  receive 


REVERBERATORY   FURNACES   WORKED  BY  HAND         51 

as  much  as  possible  of  the  radiating  heat  of  the  fire;  the  width 
differs  according  to  circumstances,  and  is  made  from  12  to  18  in. 
A  12-in.  bridge  should  be  made  entirely  of  fire-bricks,  but  if  wider 
can  be  lined  on  both  sides  with  them.  The  space  above  the 
fire-bridge  is  to  be  made  large  enough  so  that  the  hot  gases  can 
enter  the  furnace  freely  without  receiving  any  back  pressure, 
which  manifests  itself  by  flames  and  smoke  coming  out  between 
the  frame  and  door  of  the  fire-door  after  each  new  addition  of 
wood.  In  case  the  space  is  too  small  the  flame  recoils,  becomes 
short  and  overheats  the  fire-box  without  furnishing  sufficient 
heat  to  the  parts  of  the  furnace  farther  off  from  the  fire.  If  this 
space  is  large  enough  the  flame  rolls  slowly,  touching  the  roof, 
and  after  a  fresh  addition  of  wood  extends  25  to  30  ft.  into  the 
furnace.  The  fire-bridge  ought  never  to  be  built  without  air 
channels,  as  described  above. 

The  length  of  the  hearth  (see  n,  Fig.  5)  depends,  as  above 
mentioned,  entirely  on  the  nature  of  the  ore.  If  an  ore  is  rich  in 
sulphur  the  hearth  can  be  made  50  ft.  long,  but  this  is  the  limit. 
It  is  the  heat  created  by  the  combustion  of  the  sulphides  which 
makes  the  working  of  such  long  furnaces  possible.  An  ore  poor 
in  sulphur  never  could  be  heated  sufficiently  to  commence  roasting 
if  40  or  50  ft.  away  from  the  fire,  .and,  therefore,  a  large  part  of 
the  furnace  would  be  inactive,  causing  only  unnecessary  extra 
labor  to  move  the  ore.  To  insert  additional  fireplaces  does  not 
offer  any  advantages;  on  the  contrary,  it  hinders  the  execution 
of  a  delicate  chloridizing  roasting.  A  very  uneven  heating  of 
the  charge  is  caused  by  them.  Near  the  inserted  fireplace  the 
charge  gets  hot,  and  often  hotter  than  it  ought  at  that  stage  of 
roasting,  and  when  moved  to  the  next  hearth  gets  cooler  again, 
which  is  not  advantageous  to  chloridizing  roasting.  The  insertion 
of  additional  fireplaces  is  only  justified  in  mechanical  continually 
discharging  hearth  furnaces  like  the  O'Harra  and  the  Ropp  fur- 
naces, which  are  made  100  ft.  long  and  even  longer,  and  are  very 
effective  in  labor  saving  and  under  proper  conditions  do  very 
good  work;  but  their  applicability  is  confined,  like  that  of  other 
continually  discharging  mechanical  furnaces,  to  certain  kinds  of 
ore,  and  they  do  not  permit  a  really  delicate  roasting.  On  the 
first,  or  charge  hearth,  the  ore  ought  to  become  well  heated,  so 
that,  shortly  after  transferring  it  to  the  second  hearth,  blue 
flames  can  be  observed  when  the  ore  is  stirred.  Before  it  is 


52  HYDROMETALLURGY  OF  SILVER 

removed  from  here  the  oxidation  of  the  sulphides  ought  to  be 
well  started. 

Of  course,  it  takes  experience  and  skill  to  judge  the  proper 
length  to  be  given  to  the  furnace  for  a  certain  ore.  Ores  con- 
taining 20  to  22  per  cent,  sulphur,  with  considerable  iron  pyrites, 
will  roast  well  in  a  furnace  40  to  50  ft.  long,  provided  the  ore  does 
not  contain  an  excess  of  zinc  blende.  For  ores  with  about  8  per 
cent,  sulphur  a  furnace  30  ft.  long  is  sufficient,  and  ores  containing 
only  2  to  3  per  cent,  sulphur  should  not  be  roasted  in  a  long  rever- 
beratory,  but  either  in  a  Stetefeldt  or  a  Howell  furnace. 

Each  hearth  of  the  long  reverberatory  furnace  is  made  10  ft. 
long.  If  a  step  is  given  to  each  hearth,  it  serves  as  a  mark, 
but  if  the  whole  hearth  is  level,  the  points,  o,  of  the  pillars  of  the 
front  wall  mark  the  lines.  It  is  not  advisable  to  make  the  hearth 
too  wide,  thus  trying  to  increase  the  capacity  of  a  furnace.  It 
should  be  borne  in  mind  that  the  charges  have  to  be  worked  and 
moved  by  hand  labor,  and  that  extra  long  tools  are  very  hard  to 
handle;  they  tire  out  the  man,  in  consequence  of  which  the  part 
of  the  charge  next  to  the  back  wall  will  be  worked  less  than  the 
part  from  the  middle  toward  the  front.  For  the  same  reason 
the  handle  of  the  hoe  and  the  shovel  should  be  made  of  gas  pipe, 
in  order  to  make  them  as  light  as  possible.  To  the  end  of  the 
pipe  is  forged  a  solid  rod  24  to  30  in.  long,  to  which  the  hoe  or 
shovel  is  attached.  Nothing  smaller  than  a  IJ-in.  pipe  should 
be  taken,  in  order  to  make  the  handle  sufficiently  stiff  and  at 
the  same  time  easy  for  the  workman.  The  hand  has  an  easier 
hold  on  a  IJ-in.  handle  than  on  an  inch  handle. 

To  make  the  width  of  the  hearth,  measured  from  the  front  of 
the  working  door  to  the  back  wall,  10  to  11  ft.  will  be  found 
convenient,  but  it  should  never  exceed  that.  The  hight  of  the 
hearth  above  the  working  floor  should  be  2  ft.  9  in.  to  3  ft.,  so 
that  the  workman  can  throw  his  weight  on  the  handle  of  the  tool 
when  required,  which  assists  him  much. 

Of  great  importance  is  it  to  prepare  the  bottom  well  so  that 
it  remains  level  and  does  not  sink  in  different  places.  It  would 
be  too  expensive  to  build  the  whole  part  below  the  hearth  solid 
with  bricks,  for  which  reason  only  the  sides  and  ends  are  built 
solid,  while  the  inside  is  filled  with  stones,  gravel  and  sand. 
The  filling  has  to  be  done  carefully,  so  that  no  hollow  spaces  are 
left.  The  filling  is  commenced  with  coarse  rock  of  5  to  6  in., 


REVERBERATORY   FURNACES   WORKED   BY   HAND         53 

then,  after  about  8  in.  in  depth  are  filled,  finer  material  is  used 
and  washed  down  between  the  spaces  with  water.  Some  stamping 
is  to  be  recommended.  Then  another  layer  of  3  to  4  in.  of  rock  is 
added,  the  spaces  filled  with  sand  with  the  help  of  water,  and 
stamped  again.  Then  gravel,  and  as  final  layer  sand,  is  used. 
The  top  of  the  sand  layer  should  come  within  an  inch  of  the 
width  of  a  brick  if  the  upper  edge  of  the  brick  is  to  be  flush  with 
the  bottom  of  the  working-door  frame.  This  done,  an  inch 
layer  of  clay  is  spread  with  the  trowel.  A  straight-edge  and 
level  ought  to  be  used.  Then  time  has  to  be  given  for  the  clay 
to  dry,  while  work  is  being  done  on  other  parts  of  the  furnace. 
If  the  furnace  is  very  long  several  cross  walls  should  be  made, 
well  connected  with  the  side  walls. 

When  the  clay  has  sufficiently  dried,  the  brick  pavement  is 
laid.  The  hardest  bricks  are  selected  for  it.  They  are  set  on 
the  narrow  edge  and  with  their  long  side  parallel  with  the  fire- 
bridge, which  makes  the  hoe  slide  easily  over  them.  They  can 
either  be  set  dry  and  the  spaces  filled  with  sifted  sand,  or  they 
can  be  set  in  clay,  but  always  as  close  together  as  possible.  The 
hearth  next  to  the  fire  should  always  be  set  dry  in  sand  to  permit 
the  bottom  to  expand  without  bulging.  For  the  binding  rods 
which  pass  under  the  hearth,  channels  should  be  made  of  bricks, 
so  that  these  rods  will  be  kept  cool  by  air  and  can  be  easily 
inserted  or  withdrawn. 

As  regards  the  roof  r,  the  arch  over  the  hearth  should  be  made 
pretty  flat  in  order  to  spread  the  flame,  but  care  should  be  taken 
not  to  go  to  the  extreme.  A  rise  of  the  arch  of  14  to  17  in.  over  a 
10  ft.  wide  hearth  is  about  proper;  to  give  such  an  arch  only  a  rise 
of  5  to  6  in.  is  a  great  mistake  if  any  durability  of  the  furnace  is 
expected.  It  has  to  be  considered  that  in  the  heat  the  hearth 
will  expand,  and  if  the  joints  between  the  bricks  are  not  very 
close  and  carefully  made,  the  arch  will  soon  cease  to  be  an  arch, 
but  become  a  flat  plane  which  in  course  of  time  even  becomes  an 
inverted  arch,  and  soon  will  cave.  The  arch  is  usually  made 
9  in.  thick.  If  square  bricks  are  used,  which  is  generally  the 
case  in  remoter  mining  districts,  it  is  much  better  to  throw  two 
4J-in.  arches  one  on  top  of  the  other  than  to  make  only  one  arch 
by  setting  the  bricks  on  their  narrow  edge,  because  then  the 
joints  are  so  much  wider  on  the  outside  than  on  the  inside  that 
they  have  to  be  filled  with  clay  and  brick  chips.  This  filling  offers 


54  HYDROMETALLURGY  OF  SILVER 

very  little  resistance  and  makes  the  arch  weak.  The  foot  of  the 
arch  should  rest  on  the  long  sides  of  the  furnace  and  extend  also 
over  the  fireplace.  It  is  not  necessary  to  cover  the  latter  with  a 
separate  cross  arch. 

With  regard  to  the  hight  of  the  roof  above  the  hearth  bottom 
it  can  be  considered,  as  a  rule,  that  the  hight  should  be  the  greatest 
on  the  two  hearths  next  to  the  fire,  and  then  gradually  diminish 
to  the  flue  end.  In  ascertaining  this  distance  one  has  to  be 
guided  by  the  character  of  the  ore.  For  certain  ores  the  roof  of 
the  hearth  next  to  the  fire  has  to  be  made  much  higher  than  on 
the  following  hearth,  as  is  shown  in  Fig.  4,  where  the  roof  of 
this  part  of  the  furnace  is  made  5  ft.  above  the  hearth  bottom, 
or  2  ft.  higher  than  that  of  the  adjoining  hearth.  This  is  accom- 
plished by  a  step  of  2  ft.,  which,  however,  is  an  exceptional  case. 
Usually  a  long  furnace  will  do  good  work  if  the  highest  part  of 
the  arch  of  the  finishing  hearth  is  made  30  to  36  in.  above  the 
bottom  and  then  slopes  down  to  24  in.  at  the  flue  end.  This 
will  allow  live  air  to  enter  between  the  ore  and  the  fire  gases. 
The  spring  of  the  arch,  as  stated,  should  not  be  made  less  than 
14  to  17  in.  These  dimensions  are  given  for  highly  sulphureted 
ore,  and  have  to  be  varied  according  to  the  character  of  the 
ore. 

In  the  endeavor  to  reduce  the  consumption  of  fuel  to  a  mini- 
mum we  often  find  that  other  more  important  points  have  been 
sacrificed  to  this  one  object.  One  of  them  is  to  make  the  arch 
and  sides  too  low.  The  flame,  being  pressed  down  by  the  low 
arch,  comes  in  contact  with  the  ore,  and  exercises  a  reducing 
action,  which  is  adverse  to  the  principle  of  chloridizing  roasting. 
The  space  between  the  ore  and  the  roof  is  almost  completely 
filled  with  gases  from  the  combustion  of  the  fuel,  and  the  live 
air  has  no  opportunity  to  enter  deeper  into  the  furnace.  The 
little  that  enters  through  the  working  doors  is  forced  to  the  side, 
doing  some  good  work  on  its  way  to  the  flue,  but  the  main  portion 
of  the  ore  depends  for  its  oxidation  on  the  small  volume  of  unde- 
composed  air  mingled  with  the  fire  gases.  Thus  a  furnace  may 
have  plenty  of  draft  but  not  enough  air.  This  defect  in  the 
construction  is  felt  still  more  if  in  a  long  furnace  an  ore  is  to  be 
roasted  which  cannot  stand  a  high  heat  without  caking  or  suf- 
fering a  great  loss  of  silver  by  volatilization,  because  it  is  not 
possible  to  keep  the  heat  on  the  finishing  hearth  as  low  as  required 


REVERBERATORY   FURNACES  WORKED  BY  HAND         55 

and  at  the  same  time  insure  the  working  on  the  more  remote 
hearths  to  the  best  advantage. 

Instead  of  using  a  wooden  center  to  build  the  arch,  quite 
frequently  damp  sand  is  used  for  this  purpose.  The  sides  are 
finished  first,  including  the  skew  back,  then  the  center  line  is 
marked  on  the  bottom  of  the  furnace  and  wooden  sticks  set  up 
on  that  line  at  distances  of  3  to  4  ft.  and  kept  in  position  by 
some  moist  sand.  The  length  of  these  sticks  must  correspond 
with  the  hight  of  the  roof  at  their  respective  places.  All  the 
working  doors  and  other  openings  are  closed  with  boards  and 
then  the  furnace  is  filled  with  moist  sand,  or  tailings.  This  done 
the  shape  is  given  to  the  arch  with  a  straight-edge  and  trowel. 
The  finishing  is  done  with  a  thin  layer  of  lime  mortar,  on  top  of 
which  the  arch  is  built.  Clay  should  be  used  as  mortar  for  the 
arch.  After  several  days  the  sand  is  removed,  but  not  until 
buck  stays  and  binding  rods  are  placed  in  position;  otherwise, 
when  by  removing  the  sand  the  weight  and  side  pressure  of  the 
flat  arch  are  thrown  on  the  sides,  they  may  give  way  and  cause 
the  arch  to  cave. 

The  charge  hopper  h,  is  made  of  stout  sheet  iron  and  should 
be  large  enough  to  hold  a  full  furnace  charge,  which  consists  of 
1500  Ib.  to  one  ton  of  ore.  Not  to  burden  the  arch  with  this 
weight,  the  hopper  is  flanged  with  angle  iron  around  the  rim 
and  hangs  on  a  framework  which  rests  on  two  sides  of  the  furnace, 
as  shown  in  Fig.  4.  These  hoppers  are  usually  made  square,  and 
the  narrow  end,  which  is  provided  with  the  slide,  S,  corresponds 
with  a  cast-iron  extension  which  passes  through  the  arch.  This 
piece  of  casting  should  be  shaped  to  correspond  with  the  circle  of 
the  arch,  so  it  will  act  as  a  keystone.  The  hopper  is  filled  by 
means  of  dumping  cars  running  on  an  elevated  track. 

The  flue-hole  should  always  be  made  in  the  end  wall  of  the  fur- 
nace, and  not  in  the  roof,  because  the  flue  (see  e,  Fig.  4)  on  top 
of  the  furnace  is  much  in  the  way.  It  is  of  the  greatest  importance 
to  have  ample  draft  in  the  furnace.  The  draft  in  the  furnace 
depends  on  the  suction  power  of  the  stack  and  on  the  size  of  the 
flue-hole.  If  the  latter  is  made  too  small  the  gases  will  be 
throttled  and  the  furnace  will  smoke.  No  good  roasting  can  be 
performed  in  such  a  furnace;  and  as  the  draft  is  such  an  important 
factor  in  roasting,  the  flue  should  be  made  full  large  and  be 
provided  with  a  damper  to  regulate  the  draft.  Dampers  are 


56  HYDROMETALLURGY   OF   SILVER 

indispensable,  especially  if  a  number  of  furnaces  are  worked  by 
the  same  chimney.  The  furnaces  next  to  the  flue  leading  to  the 
chimney  will  receive  an  excess,  while  those  situated  farther  away 
may  not  receive  sufficient  draft. 

I  found  that  through  insufficient  supply  of  air  not  only  the 
chlorination  suffers,  but  the  loss  of  silver  also  increases.  If  the 
roof  of  a  furnace  is  of  proper  hight  to  permit  air  to  enter  between 
the  ore  and  the  gases,  and  the  fire-bridge  and  the  back  of  the 
furnace  are  provided  with  air-ducts,  the  interior  of  the  furnace  can 
be  observed,  which  makes  it  easier  to  regulate  the  draft.  The 
draft  must  be  so  regulated  that  the  fumes  evolved  are  kept  in 
motion  in  all  parts  of  the  furnace.  If  they  stagnate  around  the 
ore,  or  if  the  furnace  assumes  a  nearly  uniform  heat  throughout 
the  entire  length,  it  is  always  a  sign  of  insufficient  draft,  and  if 
the  draft  is  not  increased  the  result  will  invariably  be  a  high  loss 
of  silver  and  a  low  chlorination.  On  the  other  hand,  if  the 
flame  coming  from  the  fireplace  becomes  short  and  pointed  and 
travels  very  swiftly,  it  is  a  sign  of  too  much  draft,  which  causes 
an  unnecessary  consumption  of  fuel  and  may  cool  the  furnace  too 
much  for  it  to  do  good  work. 

If  the  flue  opening  in  the  end  wall  of  the  furnace  is  made 
4  to  5  ft.  wide,  the  sides  9  in.,  and  the  spring  of  the  arch  12  in., 
it  will  answer  for  all  kinds  of  ores. 

For  the  working  door  /,  (Fig.  5)  the  best  design  for  the 
reverberatory  furnaces  is  represented  by  Figs.  7  and  8.  It  is 
18  in.  wide  and  10  in.  high.  Its  design  is  ingenious  and  simple. 
Each  side  is  in  the  form  of  an  angle,  which  enables  the  laborer 
to  reach  through  one  door  with  his  hoe  all  points  of  a  10-ft.  hearth, 
while  if  the  two  sides  were  straight  it  would  require  two  doors 
to  each  hearth  to  reach  all  the  points.  The  top  and  bottom  plate 
of  the  frames  extending  beyond  the  sides,  and  the  angular  shape 
of  the  latter,  make  it  possible  to  cement  solidly  the  door  frame 
into  the  wall,  without  the  use  of  any  anchor-bolts,  which  are  not 
of  much  use  anyway,  as  they  invariably  work  loose  in  a  short 
time  and  then  loosen  the  bricks  around  the  door.  A  movable 
iron  plate  serves  as  door. 

Each  door  is  provided  in  front  with  a  2J-in.  roller,  the  object 
of  which  is  to  facilitate  the  working  of  the  charge.  The  long 
handle  of  the  tool  rests  on  it  and  in  stirring  it  revolves,  and  thus, 
while  taking  the  weight  of  the  tool,  lessens  the  friction.  The 


REVERBERATORY   FURNACES  WORKED  BY   HAND         57 


full  advantage,  however,  of  this  arrangement  is  only  obtained 
when  the  movement  of  the  hoe  is  at  right  angles  with  the  roller. 
As  soon  as  the  angle  is  changed  the  hoe  will  partly  slide,  or  even 
will  slide  altogether,  while  the  roller  stops.  A  far  better  contri- 
vance is  that  designed  by  G.  Kiistel,  and  shown  in  Figs.  9  A,  9  B 
and  9  C.  Instead  of  the  long  roller  he  provided  the  door  frame 


FIG.  7.  —  PLAN  OF   WORKING  DOOR. 


FIG.  8.  —  ELEVATION  OF   WORKING  DOOR. 

with  a  grooved  wheel,  which  rests  in  a  fork  movable  on  a  pivot. 
The  groove  is  large  enough  to  take  the  handle  of  the  hoe.  In 
working,  the  hoe  moves  easily  to  and  fro  on  the  wheel,  which, 
on  account  of  the  pivot  in  the  frame,  will  turn  to  any  direction 
in  obedience  to  the  hoe  and  revolve  with  the  same  ease. 

When  the  furnace  is  built,  it  has  to  be  dried  carefully  with  a 
slow  fire  for  several  days.  It  is  well  to  keep  a  small  fire  on  each 
hearth,  so  that  the  whole  furnace  will  get  about  the  same  heat. 
If  fire  is  kept  on  the  fireplace  only  it  will  take  a  very  long  time 
to  dry  such  a  long  furnace. 

In  starting  the  furnace  it  is  best  to  charge  each  hearth  through 
the  working  door  with  crude  ore.  Each  charge  will  have  to 
remain  on  the  same  hearth  until  the  charge  on  the  finishing  hearth 


58 


HYDROMETALLURGY  OF  SILVER 


is  chloridized,  which  will  be  quite  a  number  of  hours  if  the  ore  is 
highly  sulphureted,  but  after  the  first  charge  is  drawn  and  the 
others  are  moved  forward  one  hearth  each,  the  time  required 
on  the  finishing  hearth  will  be  less,  and  soon  the  whole  furnace  will 
be  in  good  working  condition.  The  discharging  is  done  through 
the  working  door  of  the  finishing  hearth,  into  wheelbarrows. 
Shortly  before  discharging  commences  the  charge  of  each  suc- 
ceeding hearth  is  raked  into  a  pile  on  the  forward  half.  As  soon 
as  the  finishing  hearth  is  discharged  the  forward  movement  of 
the  charges  begins,  and  when  done  a  new  charge  is  dropped  on 
the  first  hearth.  On  those  hearths  on  which  the  oxidizing  takes 


FIG.  9 A.  —THE  KUSTEL  WORKING 
DOOR. 

place  the  charges  should  be  worked  diligently  to  expose  the  ore 
as  much  as  possible  to  the  oxidizing  action  of  the  air.  This  not 
only  shortens  the  roasting  time  but  more  sulphuric  acid  is  formed, 
producing  a  better  sulphating  of  the  ore.  When  the  ore  becomes 
woolly,  that  is,  when  the  chloridizing  period  sets  in,  much  less 
stirring  is  required.  On  the  finishing  hearth  the  ore  should  be 
raked  to  a  thick  layer  in  the  center  of  the  hearth,  the  space  next 
to  the  fire-bridge  and  the  back  wall  being  kept  clear.  From  near 
the  working  door  the  ore  should  also  be  pushed  farther  into  the 
furnace,  to  prevent  the  cooling  of  that  portion  of  the  charge. 


REVERBERATORY   FURNACES  WORKED  BY  HAND         59 


A  thick  layer  diminishes  the  loss  of  silver  by  volatilization.  On 
this  hearth  the  ore  should  be  stirred  only  a  few  times.  If  the 
salt  is  to  be  added  in  the  furnace,  it  is  soon  found  by  observation 
and  assays  on  which  hearth  this  has  to  be  done  to  obtain  the  best 
results.  Always  consider  both  chlorination  and  volatilization. 

With  regard  to  the  fire  the  operator  has  to  be  entirely  guided 
by  the  temperature  required  on  the  finishing  hearth.  If  the  ore 
is  such  that  it  loses  much  silver  if  exposed  to  too  high  a  heat, 
which  can  be  ascertained  before  the  furnace  is  built,  then,  in 
order  not  to  reduce  the  working  capacity  of  the  furnace  by  keep- 


FIGS.  9B  and  9 C.  —DEVICE  FOR  WORKING  DOOR. 

ing  so  low  a  heat  as  the  ore  requires,  the  roof  of  the  finishing  hearth 
has  to  be  made  much  higher  above  the  hearth.  This  is  best 
done  by  dropping  that  hearth  a  step  lower  than  the  other  hearths. 
The  level  of  the  fireplace,  however,  should  not  be  dropped  too, 
but  should  be  put  in  proper  position  for  the  other  hearths, 
which,  of  course,  will  make  a  higher  fire-bridge  for  the  finishing 
hearth. 

Three  men  are  sufficient  to  attend  a  furnace  50  ft.  long,  with 
one  helper  for  each  two  furnaces.  When  a  charge  is  drawn,  two 
of  the  men  and  the  helper  wheel  the  roasted  ore  to  the  cooling 
floor,  while  the  third  man  pulls  the  charge.  The  fire  is  cared  for 
by  the  man  who  attends  to  the  two  hearths  next  to  the  fire 


60 


HYDROMETALLURGY   OF  SILVER 


on  which  the  ore  does  not  need  to  be  worked  so  frequently  as  on 
the  other  hearths.  The  removal  of  the  ashes  is  done  by  the 
helper.  The  wood  is  supplied  by  one  yard  man  to  all  the  fur- 
naces; likewise  the  hopper  is  filled  with  ore  by  one  man  for 
all  the  furnaces.  The  roasting  capacity  of  a  50-ft.  furnace 
is  about  8  to  9  tons  in  twenty-four  hours. 

THE  TWO-STORY  LONG  FURNACE 

As  a  further  improvement  in  the  hand-worked  reverberatory 
furnace  we  have  to  consider  the  two-story  furnace.  Instead  of 
building  the  furnace  in  one  direction,  say  40  ft.  long,  it  is  built 


FIG.  10.  —  LONG  REVERBERATORY  FURNACE,  TWO-STORY. 


FIG.  11. —LONG  REVERBERATORY  FURNACE,  TWO-STORY. 

in  two  stories  of  20  ft.  hearth  each.  There  is  by  this  method  not 
so  much  loss  of  heat  by  radiation.  The  hot  arch  of  the  lower 
story  warms  the  bottom  of  the  upper  hearths,  and  a  new  charge 
is  more  quickly  ignited  than  if  the  hearths  are  all  on  one  level. 
Figs.  10,  11  and  12  illustrate  the  construction  of  such  a  furnace. 


REVERBERATORY   FURNACES   WORKED   BY   HAND        61 

The  lower  two  hearths  are  11  ft.  long  each,  while  the  upper  two 
are  10  ft.  each.  A  number  of  these  furnaces  were  erected  by  me 
for  the  Mexican  Santa  Barbara  Mining  Company,  to  roast  the 
heavy  zinc-lead  ores  of  the  San  Francisco  del  Oro  mine.  In  this 
furnace  10  tons  of  ore  were  roasted  in  twenty-four  hours.  The 
oxidizing  was  done  on  the  upper  two  hearths.  Through  an  open- 
ing in  the  bottom  of  the  second  upper  hearth  the  charge  was 


18 'Flue 


FIG.  12.  —  SECTION  THROUGH  A  B 
(Fie.  10). 

dumped  into  the  first  hearth  of  the  lower  furnace.  The  salt  was 
added  in  the  upper  hearth  just  before  dropping  the  charge  into 
the  lower  hearth.  In  this  way  the  salt  was  well  mixed  with  the 
ore.  The  construction  of  these  two-story  furnaces  permits  the 
insertion  of  an  auxiliary  fireplace  across  the  width  of  the  hearth, 
which  can  be  used  to  advantage,  because  the  heat  is  spread  over 
the  whole  width  of  the  hearth,  which  is  not  the  case  in  a  long 
furnace,  where,  if  an  auxiliary  is  used,  the  fire  ha&  to  enter  the  side 
of  the  furnace.  In  roasting  the  San  Francisco  del  Oro  ore  the 
auxiliary  fire  was  used  only  for  a  short  time  after  a  new  ore 
charge  entered  the  furnace,  for  the  purpose  of  reducing  the  cooling 
effect  of  the  cold  charge  on  the  other  charge,  which  was  in  the 
state  of  roasting,  and  to  ignite  the  fresh  charge  more  quickly. 
The  fire  was  stopped  as  soon  as  the  fresh  charge  showed  the 
sulphur  flame.  The  gases  from  the  lower  furnace  entered  the 
upper  furnace  through  two  flues,  one  on  each  side  of  the  fire- 
bridge proper,  so  the  rising  gases  did  not  interfere  with  the  flame 
of  the  auxiliary  fire. 

At  Sombrerete,  Mexico,  when  it  became  necessary  to  rebuild 
two  of  the  40-ft.  furnaces,  I  changed  them  into  20-ft.,  two-story 
furnaces,  and  increased  thereby  the  working  capacity  of  each  by 
nearly  two  tons  per  day. 


VIII 

MECHANICAL  ROASTING  FURNACES 

IN  this  chapter  will  be  included  only  such  mechanical  roasting 
furnaces  as  are  specially  adapted  for  chloridizing  roasting.  There 
are  two  classes  of  such  furnaces:  one  in  which  the  ore  is  roasted 
in  charges;  and  the  second  in  which  the  roasting  is  continuous, 
that  is,  furnaces  in  which  at  one  end  a  continuous  stream  of 
raw  ore  enters,  while  at  the  other  a  continuous  stream  of  roasted 
ore  leaves  the  furnace. 

As  stated  above,  in  no  mechanical  furnace  can  the  process 
of  roasting  in  all  its  stages  be  so  well  controlled  as  in  a  rever- 
beratory  furnace  worked  by  hand,  and,  therefore,  their  applica- 
bility is  much  more  limited  to  certain  classes  of  ore.  However, 
if  applied  to  suitable  ores,  they  do  very  good  work,  and,  where 
labor  is  expensive,  are  more  economical.  In  Mexico,  where  labor 
is  cheap,  the  mechanical  furnaces  proved  successful  only  in 
exceptional  cases.  All  mechanical  furnaces  are  connected  with 
more  or  less  machinery  and  require  frequent  replacement  of  the 
wearing  parts.  These  parts  of  machinery  and  castings  are  sent 
from  the  United  States,  and  are  rather  costly  by  the  time  they 
land  in  some  remote  mining  place  in  the  mountains.  This, 
however,  would  not  be  so  important  a  factor  if  it  were  not  for 
other  inconveniences  connected  with  it,  as  the  long  time  it  takes 
to  get  these  parts,  and,  perhaps  the  sudden  breakdown  of  parts 
of  which  duplicates  may  not  be  on  hand,  requiring  quite  a  long 
shutdown  of  one  of  the  furnaces,  which  always  represents  a  large 
percentage  of  the  roasting  capacity  and  with  it  of  the  producing 
capacity  of  the  works. 

The  erection  of  mechanical  furnaces  for  ores  for  which  they 
were  not  suitable  has  caused  many  serious  failures. 


62 


MECHANICAL   ROASTING   FURNACES 


63 


(1)  MECHANICAL  FURNACES  FED  BY  CHARGES 

(a)  The  Bruckner  Revolving  Furnace.  —  This  ingenious  device 
of  a  chloridizing  furnace  was  successfully  introduced  by  its 
inventor,  Mr.  Bruckner,  in  Colorado,  in  1867.  The  furnace  con- 
sists, as  illustrated  by  Figs.  13  and  14,  of  a  cylinder  of  boiler  iron, 
the  ends  of  which  are  closed  save  a  circular  central  opening  on 
each  end.  Two  circular  tracks  are  fastened  around  the  cylinder 
placed  at  even  distances  from  the  ends.  With  these  tracks  the 
cylinder  rests  on  four  strong  wheels.  A  revolving  motion  is 
imparted  to  the  cylinder  either  by  friction,  in  which  case  one  of 


FIGS.  13  and  14.  —  BRUCKNER  ROASTER. 

the  four  wheels  is  made  to  revolve,  or  the  motion  is  imparted  by 
pinion  and  cogs,  in  which  case  a  cast-iron  ring  with  cogs  is  fas- 
tened to  the  cylinder.  At  one  end  of  the  cylinder  is  placed  a 
fire-box,  the  throat  of  which  corresponds  with  the  central  end 
opening  of  the  cylinder,  while  the  opposite  opening  corresponds 
with  a  circular  hole  in  the  flue.  Four  doors  are  placed  diamet- 
rically opposite,  two  on  each  side.  These  doors  serve  for  charging 
and  discharging  the  furnace.  Above  the  furnace  is  placed  a 
hopper  large  enough  to  hold  a  charge  of  ore.  The  hopper  has 
two  outlet  spouts,  each  provided  with  a  slide,  which  correspond 
with  the  furnace  doors.  The  shell  as  well  as  the  ends  are  lined 


64  HYDROMETALLURGY   OF  SILVER 

with  bricks.  Provision  is  made  in  the  driving  mechanism  to 
regulate  the  speed  from  one  revolution  in  one  minute  to  one 
revolution  in  three  minutes.  Some  of  the  furnaces  are  so  con- 
structed that  the  two  ends  of  the  cylinder  are  slightly  contracted 
in  order  to  facilitate  the  discharging  of  the  ore,  but  this  compli- 
cates the  lining  and  reduces  the  capacity,  and  is  actually  not 
necessary,  as  a  straight  cylinder  discharges  very  nicely,  and  if 
a  little  of  the  charge  does  remain  in  the  furnace,  it  helps  to  heat 
the  new  charge. 

The  manipulations  of  this  furnace  are  as  follows :  The  two 
doors  of  one  side  are  opened  and  the  furnace  revolved  until  the 
doors  come  right  under  the  two  spouts  of  the  hopper,  when  the 
two  slides  are  withdrawn  and  the  charge  allowed  to  drop  into 
the  furnace.  Then  a  quarter  turn  is  given  to  bring  the  doors  in 
proper  position  for  the  man  to  close  them.  They  are  fastened  as 
tight  as  possible  by  means  of  a  wedge.  Though  the  lid  and  the 
flange  of  the  cast-iron  door  frame  are  faced  it  is  not  possible, 
especially  if  the  furnace  has  been  in  use  for  some  time,  to  close  the 
door  perfectly  tight,  and  when  the  furnace  revolves,  some  ore 
will  leak  out  through  each  door.  This  takes  place  only  as  long 
as  the  charge  is  crude  and  stops  when  actual  roasting  is  in  progress. 
This  leakage  is  the  more  annoying  as  the  ore  is  crude  and  mixes 
with  roasted  ore  underneath  the  furnace.  To  prevent  this  leak- 
age the  joint  of  lid  and  flange  should  be  plastered  from  the  out- 
side with  clay,  or  better  with  paste  of  sifted  wood  ashes  and 
salt.  Before  charging  the  furnace  should  be  well  heated,  and 
during  charging  the  draft  checked  so  that  not  too  much  dust  is 
carried  into  the  flue. 

In  the  beginning  a  strong  fire  is  kept,  but  as  soon  as  the  sul- 
phides are  well  ignited  the  fire  is  allowed  to  go  out.  The  heat  de- 
veloped by  the  oxidation  is  considerable,  and  if  further  increased 
by  the  fire  a  caking  and  balling  of  the  ore  would  take  place. 
These  furnaces  are  usually  6  ft.  in  diameter  inside  the  lining  and 
16  ft.  long,  and  take  a  charge  of  5  to  5J  tons  of  ore.  Usually  two 
charges  can  be  roasted  in  twenty-four  hours.  About  three  to 
four  hours  of  each  charge  the  furnace  can  run  without  fire,  then 
the  salt  is  added  and  the  roasting  continued  with  a  moderate 
fire.  A  hole  back  of  the  flue  permits  the  observation  of  the 
temperature  and  the  taking  of  samples  with  a  long-handled  scoop. 
During  chloridizing  the  charge  in  the  furnace  assumes  an  inclined 


MECHANICAL   ROASTING   FURNACES  65 

position  up  to  45  deg.,  the  weight  of  which  acts  against  the  direc- 
tion of  the  motion,  and  if  the  clutch  is  thrown  out  in  order  to 
stop  the  furnace,  this  weight  will  pull  the  furnace  back  nearly  a 
quarter  of  a  turn.  To  open  the  doors  to  add  the  salt  or  to  dis- 
charge the  furnace,  it  is  necessary  that  the  furnace  should  be 
stopped  at  a  position  convenient  to  the  roaster  man.  In  order 
to  accomplish  this  the  end  of  an  iron  bar  is  pressed  between  the 
track  and  the  wheel  at  the  side  at  which  the  movement  of  both 
is  outward.  When  the  doors  are  in  the  right  position  the  clutch 
is  thrown  out,  but  the  furnace  is  prevented  from  revolving  back 
by  the  iron  bar,  which  becomes  clamped  in  very  tightly.  By  means 
of  a  short-handle  shovel  the  salt  is  introduced  through  the  two 
doors  and  well  scattered  over  the  whole  surface  of  the  ore.  The 
salt  decrepitates  violently  and  in  this  way  becomes  very  evenly 
divided.  After  the  salt  is  added  and  the  furnace  revolves  again, 
the  ore  becomes  woolly,  as  in  the  reverberatory,  and  assumes  a 
still  more  upright  position. 

If  the  ore  is  sufficiently  sulphureted  and  the  salt  is  added  it 
will  not  be  necessary  to  start  the  fire  again.  There  is  enough 
heat  stored  in  the  charge  to  finish  the  process  of  chlorination 
(see  remarks  on  chloridizing  self-roasting),  in  which  case  the  con- 
sumption of  fuel  is  very  small.  This  mode  of  roasting  is  only 
admissible  if  the  ore  is  to  be  roasted  for  lixiviation.  For  amalga- 
mation a  second  fire  is  indispensable,  because  the  base-metal 
chlorides  have  to  be  decomposed  or  volatilized. 

When  the  charge  is  finished  two  cars  are  pushed  under  the 
furnace,  one  for  each  door.  They  are  large  enough  to  receive  the 
full  charge  of  the  furnace.  All  four  doors  are  opened,  the  lids 
kept  in  position  by  a  proper  contrivance,  and  the  furnace  is  re- 
volved again.  While  the  furnace  is  prepared  for  discharging,  a 
good  strong  fire  should  be  started  again,  to  heat  it  for  the  next 
charge.  The  receiving  cars  are  made  narrow  and  long,  so  that 
no  ore  is  dropped  beyond  the  rim  of  the  car,  because,  especially 
in  the  beginning,  the  ore  will  pass  through  the  door  over  a 
large  arc. 

The  cooling  floor  is  situated  5  or  6  ft.  lower  than  the  track 
under  the  furnace,  which  track  extends  some  distance  on  iron 
trestle  over  the  cooling  floor.  The  body  of  the  cars  is  shaped 
like  a  hopper  with  bottom  discharge,  and  closed  by  a  slide  which 
is  worked  by  a  lever.  As  each  car  holds  about  2J  tons  and  is 


66  HYDROMETALLURGY  OF  SILVER 

hot,  they  are  pulled  out  from  under  the  furnace  by  means  of  a 
windlass  and  chain. 

The  revolving  motion  of  the  Bruckner  furnace  should  be 
slow.  There  is  not  the  least  advantage  in  whirling  the  ore  around 
and  around  in  the  furnace.  When  new  ore  particles  are  brought 
to  the  surface  it  takes  some  time  to  undergo  oxidation;  why  not, 
then,  give  them  the  required  time  to  be  in  contact  with  the  air  be- 
fore again  immersing  them  under  the  surface?  I  found  in  some 
works  large  Bruckner  furnaces  set  to  make  two  and  even  two  and 
a  half  revolutions  per  minute,  whereas  a  speed  of  one  revolution  in 
three  minutes  is  ample.  A  Bruckner  furnace,  when  charged, 
weighs  about  16  tons;  this  divided  on  four  wheels  makes  4  tons 
to  the  wheel,  a  rather  heavy  weight,  especially  if  we  consider 
that  the  whole  of  this  weight  is  pressing  against  the  space  of  con- 
tact between  the  ring-track  of  the  furnace  and  the  surface  of  the 
wheel,  which  is  very  small.  The  effect  of  this  high  pressure  is 
shown  by  the  way  the  wheels  and  the  ring-tracks  wear.  When 
the  furnace  is  revolving  it  can  be  observed  that  continually  thin 
scales  of  iron  up  to  the  size  of  a  finger  nail  are  dropping  from  the 
face  of  the  wheels  and  tracks.  Now  if  two  revolutions  are  made 
in  one  minute,  instead  of  one  revolution  in  three  minutes,  the 
wear  will  be  six  times  as  great,  to  say  nothing  about  the  greater 
power  which  is  required  for  the  faster  motion.  If  the  capacity 
of  the  furnace  should  be  increased  by  it  six  times  it  would  be 
different,  but  this  is  not  the  case;  a  charge  roasted  at  a  high  fur- 
nace speed  takes  just  as  long  to  be  finished  as  a  charge  does  when 
rotated  at  a  moderately  slow  speed.  Besides  much  increasing 
the  wear,  a  high  speed  can  cause  serious  trouble  if  the  ore  has  a 
tendency  to  ball,  in  which  case  it  may  happen  that  the  whole 
charge  is  transformed  into  balls  from  the  size  of  a  pea  to  that  of 
a  cocoanut,  without  leaving  any  fine  stuff  at  all. 

One  of  the  main  advantages  of  the  Bruckner  furnace  is  the 
fact  that  the  ore  in  this  furnace  is  kept  in  a  thick  layer  and  still 
permits  a  thorough  oxidation  and  chlorination.  As  explained 
above,  the  loss  of  silver  by  volatilization  is  much  less  if  the  ore 
during  roasting  can  be  kept  in  a  thick  layer,  because  a  large 
portion  of  the  evolving  volatile  chlorides  is  kept  back  in  the  ore 
as  by  a  filter,  and  consequently  much  less  silver  will  be  carried 
away. 

Under  equal  conditions,  it  will  be  found  that  an  ore  loses  less 


MECHANICAL  ROASTING  FURNACES  67 

silver  if  chloridized  in  a  Bruckner  than  in  any  other  furnace  in 
which  the  ore  is  roasted  in  a  thin  layer. 

Ores  not  suitable  for  the  Bruckner  furnace  are  those  which, 
on  account  of  a  large  percentage  of  lead,  antimony,  etc.,  cake 
easily,  because  by  the  continuous  rolling  of  the  ore  any  lumps 
which  may  have  formed  cannot  be  mashed  shortly  after  they 
are  formed,  but  increase  much  in  size  by  the  continual  rolling, 
which  makes  them  hard  and  dense. 

(6)  The  0.  Hofmann  Improved  Bruckner  Furnace.  —  In  the 
Bruckner  furnace  the  ore  is  exposed  to  a  rather  uneven  heat. 
The  part  next  to  the  fire  receives  always  the  highest  heat,  while 
that  at  the  other  end,  16  ft.  away,  will  receive  much  less,  in 
some  cases  not  even  enough  without  overheating  the  fire  end. 
This  is  a  defect  of  the  furnace  which  is  of  no  consequence  if  a 
highly  sulphureted  ore  is  treated,  which  can  be  subjected  to 
self-roasting,  because  it  creates  ample  heat  by  itself  to  become 
uniformly  hot  through  the  entire  length  of  the  furnace;  but  it  is 
a  defect  much  felt  if  ore  poor  in  sulphur  is  to  be  roasted,  which 
has  to  receive  nearly  all  the  required  heat  from  the  fire,  or  an  ore 
which  has  to  be  roasted  at  a  low  heat,  because  it  is  apt  to  lose 
much  silver  by  volatilization,  or  lumps  easily  at  a  higher  tem- 
perature (see  chloridizing  roasting  with  steam,  page  34).  Mr. 
Bruckner  was  aware  of  this  defect,  and  he  tried  to  remedy  it  by 
inserting  into  the  furnace  a  diaphragm  made  of  cast-iron  pipes. 
This  diaphragm  was  set  at  an  angle  of  about  15  deg.  to  the  axis. 
It  had  a  diagonal  position  extending  through  the  whole  length 
of  the  furnace  and  was  intended  to  move  the  ore  from  one  end 
to  the  other  and  back,  in  order  to  produce  a  uniform  heating  of 
the  charge.  This  device,  however,  did  not  give  the  satisfaction 
expected  and  was  soon  abandoned,  especially  on  account  of  the 
inconvenience  the  diaphragm  caused  in  cleaning  the  furnace  from 
the  crust,  which  has  to  be  done  from  time  to  time,  and  of  the 
short  life  of  the  pipes,  though  they  projected  through  the  shell 
of  the  furnace  to  permit  air  to  pass  through. 

Confronted  with  the  necessity  of  obviating  this  defect  of  the 
Bruckner  furnace,  because  the  rich  ores  of  the  Silver  King  mine, 
Arizona,  could  not  be  roasted  successfully  in  the  Bruckner 
owing  to  this  defect,  I  changed  the  arrangement  of  the  fur- 
naces, inasmuch  as  I  attached  a  fire-box  and  flue  arrangement 
to  each  end  of  the  furnace,  which  enabled  me  to  heat  either 


68 


HYDROMETALLURGY  OF  SILVER 


end  of  the  furnace  by  changing  at  intervals  the  course  of  the 
flame. 

Figures  15,  16,  and  17  represent  the  arrangement.  Between 
the  fireplace  proper  and  the  furnace  is  situated  the  flue,  extending 
downward  to  the  dust-chambers,  Fig.  15.  This  flue  is  provided 
with  a  damper.  The  other  end  of  the  furnace  is  provided  with 
exactly  the  same  arrangement.  The  dust-chambers  from  both 
ends  are  connected  with  the  main  flue  leading  to  the  stack. 
Before  they  connect  there  is  an  additional  damper,  Fig.  16,  on 
each  side,  to  make  it  sure  that,  if  the  dampers  of  one  side  are 
closed,  no  draft  passes  through  on  that  side. 


=*n 


iu-\-m 
•  ra1 1 


j  1 ...-.  • _ 


3H- 

BP.J..  i          ~n 


FIGS.  15-17.  —  HOFMANN  IMPROVED  BRUCKNER  FURNACE. 


If  the  fire  is  kept  at  one  side  for  some  time  the  dampers  of 
that  same  side  are  opened  and  those  of  the  other  are  closed, 
and  the  fire  started  there.  The  flame  now  traverses  the  furnace 
in  the  other  direction,  and  the  ore  at  that  end  will  be  exposed  to 
the  same  heat  as  the  other  was  before.  This  changing  can  be 
done  at  intervals  to  suit  the  character  of  the  ore.  The  changing 
of  the  fire  does  not  cause  any  trouble,  as  the  opposite  fireplace 
is  still  warm  enough  to  ignite  the  wood  when  the  change  is  made. 

The  results  obtained  with  this  furnace  have  been  very  satis- 


MECHANICAL   ROASTING  FURNACES  69 

factory.  I  obtained  with  it  good  results  in  roasting  the  ores  of 
the  North  Mexican  Silver  Mining  Company,  Mexico,  which  were 
very  poor  in  sulphur  and  could  not  be  heated  sufficiently  at 
the  opposite  end  with  a  fire  at  one  end  only.  With  the  double 
arrangement  quite  satisfactory  results  were  obtained. 

If  a  number  of  these  furnaces  are  built  in  a  row  the  fires  of 
all  of  them  have  to  be  changed  at  the  same  time,  which  is  neces- 
sary on  account  of  the  dampers  in  the  two  wings  of  dust-chambers 
and  on  account  of  the  man  attending  the  furnaces.  If  the  fires 
are  all  on  one  side  he  will  and  can  attend  better  and  to  more 
furnaces  than  if  the  fires  are  partly  on  one  and  partly  on  the 
other  side. 

There  is  quite  a  serious  omission  in  the  construction  of  the 
regular  Bruckner  furnace,  and  that  is  that  there  is  no  provision 
made  for  the  admittance  of  live  air  into  the  furnace.  Air  being 
such  an  important  factor  in  roasting  it  is  absolutely  necessary 
to  have  some  means  to  admit  air  if  the  roasting  is  to  be  conducted 
intelligently. 

To  keep  the  fire-door  open  for  this  purpose  is  not  to  be  recom- 
mended, as  this  shortens  the  flame.  In  the  Hofmann  furnace 
provision  is  made  for  an  air  inlet  by  making  the  circular  cast-iron 
throat  of  the  fire-box  longer  than  usual  and  by  leaving  in  the 
lower  half  of  the  same  a  sufficiently  large  opening  for  the  air. 
The  size  of  the  opening  can  be  regulated  by  a  hinged  door  and  a 
lever. 

On  the  periphery  of  the  furnace  and  near  each  end  is  a  small 
door  for  taking  samples.  These  small  doors  are  easy  to  handle, 
and  they  can  be  opened  and  closed  and  a  sample  taken  without 
stopping  the  furnace,  which  is  quite  convenient. 

The  spouts  of  the  hopper  have  to  be  made  to  stand  pretty 
high  above  the  furnace  in  order  to  permit  the  passage  underneath 
of  the  door  and  the  eye  to  which  the  door  is  keyed  when  open 
for  discharging.  This  causes  considerable  spilling  of  ore  during 
charging.  To  prevent  this,  the  spouts  are  each  provided  with  a 
sliding  sleeve  kept  in  position  by  a  lever  and  weight,  Fig.  17. 
When  the  furnace  is  stopped  with  the  doors  open  for  charging, 
the  weight  of  the  lever  is  removed  and  the  sleeve  lowered  until 
it  projects  into  the  door.  Then  the  hopper  slide  is  pulled.  This 
arrangement  permits  very  clean  work. 

The  lining  of  the  furnace  has  to  be  done  very  carefully.     The 


70  HYDROMETALLURGY  OF  SILVER 

door  frames  projecting  inside  should  conform  with  the  circle  of 
the  lining.  Specially  made  bricks  to  fit  the  circle  of  the  arch 
should  be  used  only;  if  not,  the  lining  will  soon  come  out.  To 
facilitate  the  work  of  lining  it  is  well  to  rivet  to  the  steel  shell  of 
the  furnace  six  angle-iron  ribs  the  whole  length  of  the  cylinder, 
which  will  divide  it  into  six  equal  sections.  The  projecting  part 
of  the  rib  should  be  3^  in.,  so  that  when  the  bricks  are  laid  the 
latter  project  half  an  inch  above  the  rib.  The  groove  which  is 
formed  by  it  is  filled  with  clay.  Thus  each  section  forms  an  arch 
for  itself,  the  angles  serving  as  skew-backs.  Each  should  be 
well  keyed.  In  lining,  the  bricklayer  can  then  bring  the  cylin- 
der always  in  the  most  convenient  position  for  his  work,  as  there 
is  no  danger  of  caving  even  if  that  part  of  the  lining  which  was 
made  first  comes  to  stand  right  above  him.  If,  in  course  of  time, 
from  some  reason  or  other,  part  of  the  lining  should  become 
defective,  the  bricks  of  the  bad  section  can  be  taken  out  and 
replaced  without  disturbing  the  other  parts  of  the  lining.  The 
ends  should  be  lined  first,  so  that  the  lining  of  the  cylinder  will 
abut  against  it. 

As  the  furnaces  of  this  type  when  charged  are  very  heavy, 
the  tracks  with  which  they  rest  on  the  wheels  should  be  made 
very  solid,  allowing  considerable  iron  for  wear.  At  the  places 
at  which  the  two  tracks  are  to  come  a  wrought-iron  band  an 
inch  thick  and  as  wide  as  the  track,  including  its  two-flange  pro- 
jection, is  to  be  strongly  riveted  to  the  shell.  To  this  band  the 
track  is  fastened  by  means  of  tap-screws.  Bolts  will  not  answer 
because,  if  the  track  is  worn  and  has  to  be  replaced  by  a  new  one, 
it  will  be  found  that  many  of  the  nuts  are  so  tightly  roasted  in 
that  the  bolts  will  be  turned  off  in  trying  to  unscrew  the  nuts, 
and  to  renew  these  bolts  means  the  taking  out  of  a  part  of  the 
lining,  which  ought  not  to  be.  If  a  tap-bolt  breaks,  the  piece  in 
the  wrought-iron  band  can  be  bored  out  and  a  new  tap-bolt  put 
in  without  difficulty.  It  is  best  to  have  the  track  cast  in  about 
four  segments.  The  joints  of  the  segments  should  not  be  square 
with  the  track,  but  slanting,  because  the  furnace  will  revolve 
more  smoothly  and  the  chipping  of  the  edges  of  the  joints  will  be 
much  less. 

As  in  any  other  roasting  furnace,  in  course  of  time  a  crust  is 
formed,  which  has  to  be  removed  from  time  to  time.  This  is 
usually  done  by  cooling  down  the  furnace  and  having  men  remove 


MECHANICAL   ROASTING   FURNACES  71 

the  crust  by  picks,  shovels  and  other  tools.  When  cold  the 
crust  is  quite  hard,  and  the  removal  of  it  endangers  and  weakens 
the  lining.  While  hot  this  crust,  however,  is  rather  soft,  which 
suggested  to  me  a  method  by  which  the  furnace  can  be  quickly 
cleaned  without  cooling  or  shutting  down  the  furnace.  When 
the  furnace  needs  cleaning,  after  it  has  been  discharged,  a 
charge  is  made  up  of  bricks,  fire-bricks  if  possible,  and  the  furnace 
revolved  while  a  very  strong  fire  is  kept.  The  heat  softens  the 
crust,  and  the  bricks  in  moving  in  the  furnace  gradually  shave 
off  the  crust  without  injuring  the  lining.  This  is  done  in  a  few 
hours,  and  when  the  bricks  and  crust  are  discharged  the  furnace 
is  not  only  in  shape  but  also  heated  to  receive  a  new  charge. 
Besides  this,  the  men  are  relieved  from  an  unhealthful  job. 

There  are  other  roasting  furnaces  which  properly  belong  to 
the  class  of  mechanical  furnaces  fed  by  charge,  like  the  revolving- 
hearth  furnaces;  but  they  never  were  used  much  in  actual  prac- 
tice, they  are  not  very  convenient,  and  are  of  small  capacity,  so 
that  I  can  without  hesitation  leave  them  undescribed. 


(2)  MECHANICAL  ROASTING  FURNACES  WITH 
CONTINUOUS  FEEDING 

To  this  class  of  furnaces  belong  the  O'Harra,  the  Ropp  the 
Howell-White  and  the  Stetefeldt.  The  Brown  and  Pearce  fur- 
naces, while  they  are  excellent  for  oxidizing  roasting,  are  not 
quite  suitable  for  chloridizing  roasting. 

(a)  The  O'Harra  Furnace. — This  furnace  was  the  first  mechani- 
cal furnace  with  continuous  feeding  devised  on  the  Pacific  slope. 
It  is  very  ingeniously  arranged  and  is,  in  improved  form,  still  in 
use.  In  the  following  I  give  a  description  of  this  furnace  by  the 
late  G.  Kiistel,  who  had  much  experience  with  it: 

"This  furnace  was  first  tried  in  1862  or  1863  in  Dayton, 
Nevada,  and  later  three  of  them  were  built  in  Flint,  Idaho.  The 
main  feature  of  this  furnace  is  the  endless  chain  to  which  two 
oval  rings  are  attached,  the  wings  being  as  wide  as  the  cross- 
section  of  the  hearth.  To  these  rings  are  fastened  the  plows  or 
shoes  by  which  the  ore  is  gradually  pushed  forward.  The  hearth 
of  the  furnaces  built  in  Flint  were  104  ft.  long  and  nearly  5  ft. 
wide.  Eighty  feet  of  this  hearth  were  covered  by  an  arch  12  in. 
high;  attached  to  it  were  three  fireplaces  —  two  on  one  side,  and 


72  HYDROMETALLURGY  OF  SILVER 

one  between  the  two  on  the  other  side.  At  one  end  was  the  feed- 
ing hearth,  which  was  not  covered  by  the  arch,  and  on  which  the 
ore  was  continually  delivered  from  the  stamp  battery  by  mechani- 
cal contrivances.  The  motion  of  the  ore  was  effected  by  an  endless 
chain,  passing  over  two  chain  wheels,  one  at  each  end.  To  this 
chain  two  oblong  flat  rings  were  attached,  each  provided  with 
eight  shovels  or  plows,  so  arranged  that  while  one  of  the  rings 
shoveled  the  ore  toward  the  center  line,  the  other  pushed  it  back 
again  toward  the  sides  every  three  or  four  minutes  (or  in  shorter 
intervals  if  more  ore  is  charged).  The  ore  not  only  changed  its 
place  to  the  right  and  left,  but  it  also  moved  forward  by  degrees, 
so  that  in  course  of  six  hours  from  the  beginning  it  commenced 
to  be  discharged,  passing  18  ft.  over  a  cooling  hearth.  On  both 
ends  of  the  furnace  were  iron  doors  hung  on  hinges  which  were 
opened  by  the  ring  every  time  it  passed. 

"The  whole  plant  at  Flint  was  arranged  to  work  automatically. 
The  five  batteries,  of  five  stamps  each,  had  on  both  long  sides  end- 
less screws,  by  which  the  crushed  ore  was  forwarded  in  proportion 
as  it  discharged  from  the  battery,  and  dropped  into  an  eleva- 
tor. Having  been  lifted  about  15  ft.,  it  was  conveyed  again  by 
endless  screws  along  the  feeding  hearths  of  all  three  furnaces. 
The  discharge  of  this  conveyor  was  so  regulated  that  each  feeding 
hearth  received  an  even  part  of  the  ore.  The  ore  mixed  with 
5  per  cent,  of  salt  was  spread  on  iron  plates  behind  the  batteries 
(heated  by  the  hot  gases  from  the  furnaces,  conveyed  through 
the  flue  and  under  the  plates).  When  charged  into  the  battery, 
the  ore  was  not  further  handled  till  it  came  out  of  the  furnace 
perfectly  roasted. 

"There  is  only  one  obstacle  connected  with  this  mechanical 
furnace.  The  shoes  or  shovels,  touching  the  sides  of  the  furnace, 
wear  off  by  degrees,  leaving  a  space  which  is  taken  up  by  the  ore. 
This  part  of  the  ore  along  the  wall  hardens  and  increases  in 
amount  in  the  furnace  till  new  shoes  are  put  in.  By  these  the 
crust  of  one-half  to  three-quarters  of  an  inch  thick  is  broken  off 
and  carried  out.  From  the  Rising  Star  ore  these  crusts  contain 
nearly  as  much  silver  chloride  as  the  well-roasted  ore;  they  are, 
nevertheless,  disagreeable,  but  some  means  might  be  devised  by 
which  this  inconvenience  could  be  avoided. 

"The  ore  from  the  Rising  Star  mine  at  Flint  contained 
argentiferous  fahlerz,  miargyrite,  ruby  silver,  zinc  blende,  galena, 


MECHANICAL  ROASTING  FURNACES 


73 


iron  pyrites,  and  sulphide  of  antimony.  At  an  average  the  ore 
contained  between  69  to  77  ounces  silver  per  ton  and  some  gold. 
The  gangue  was  quartz.  It  was  crushed  through  sieves  with 
forty  holes  to  the  inch,  together  with  5  per  cent,  of  salt.  The 
ore  in  the  furnace,  when  reaching  the  first  fireplace,  commenced 
to  roast  oxidizingly.  Between  this  fireplace  and  the  second,  which 
was  on  the  other  side,  the  chlorination  began  at  an  increased 
heat.  Between  the  second  and  third  fireplace  the  chlorination 
was  finished  at  a  high  red  heat.  Although  not  more  than  5  per 
cent,  of  salt  was  added,  the  roasted  ore  contained  about  90  per  cent, 
of  the  silver  converted  into  a  chloride.  The  gases,  containing 
free  chlorine  and  chloride  combinations  emitting  chlorine,  coming 
in  contact  with  the  surface  of  the  ore  while  passing  over  it  for  a 


FIG.  18.  — O'HARRA  FURNACE. 

space  of  eighty  feet,  have  a  chloridizing  influence  on  it,  replacing 
thus  a  certain  amount  of  salt. 

"The  capacity  of  the  three  furnaces  was  more  than  twenty 
tons.  Each  one  could  easily  treat  ten  tons  of  the  Rising  Star 
ore  in  twenty-four  hours.  The  roasted  ore  was  treated  by 
amalgamation  in  pans." 

Mr.  Kiistel  continues: 

"O'Harra's  furnace  is  now  greatly  improved  (Fig.  18).  It  is 
built  in  two  stories,  so  that  when  the  chain  comes  out  of  the 
lower  hearth  it  turns  into  that  of  the  upper  story.  The  chain  is 
heavy  and  there  is  no  trouble  whatever.  Although  the  chain 
in  its  course  through  the  red-hot  furnace  is  exposed  to  red  heat, 


74  HYDROMETALLURGY  OF  SILVER 

it  nevertheless  does  not  become  so  hot  as  to  suffer  any  injury 
from  it;  there  is  a  wooden  framework  on  either  end  of  the  furnace 
over  which  the  chain  and  plows  move  in  the  open  air,  which 
prevents  them  from  getting  too  hot.  The  furnace,  of  the  latest 
construction,  is  eight  feet  wide  and  from  forty  to  a  hundred  feet 
long,  with  four  fires,  two  on  each  side,  directly  opposite  each 
other.  The  first  two  fires,  where  the  ore  comes  in  contact  there- 
with, are  divided,  so  that  one-half  of  the  flame  goes  direct  to  the 
lower,  and  the  other  half  to  the  upper  hearth,  through  an  opening 
in  the  arch  of  the  fire-chamber.  There  is  a  fire-clay  damper  to 
regulate  the  flame.  The  other  two  fires  are  opposite  each  other, 
so  that  the  heat  is  uniform  over  the  whole  hearth.  These  last 
two  fires  are  regulated  by  the  ash-pit  dampers. 

"The  endless  chain  has  two  triangular  frames,  with  plow- 
shoes  on  each  side.  The  cooling  space  is  built  in  proportion  to 
the  length  of  the  hearth.  A  furnace  of  this  kind,  fifty  feet  long 
and  eight  feet  wide,  can  roast  from  thirty  to  forty  tons  in  twenty- 
four  hours,  with  the  help  of  only  two  men,  consuming  about  two 
cords  and  a  half  of  wood.  The  chlorination  runs  up  to  90  and 
95  per  cent. 

"The  working  of  this  furnace  is  not  expensive,  as  two  men, 
one  at  daytime  the  other  at  night,  can  attend  the  roasting  of 
forty  tons. 

"A  remarkable  feature  of  O'Harra's  furnace  is  the  very  small 
amount  of  dust  that  is  carried  off  by  the  draft.  Another  pecu- 
liarity of  the  furnace  is  for  drying  ore  in  pieces  the  size  of  a  man's 
fist.  One  furnace  near  Shasta,  California,  dries  40  tons  of  ore  in 
twenty-four  hours,  at  a  small  expense. " 

The  construction  of  this  furnace  does  not  offer  many  facilities 
for  regulating  the  process,  and  therefore  it  can  be  used  for  chlo- 
ridizing  roasting  only  for  a  particular  class  of  ore.  To  roast  in 
this  furnace  ores  which  are  apt  to  lose  much  silver  by  volatilizing 
would  be  rather  risky.  Moreover  this  furnace  is  not  suitable  for 
roasting  ores  which  require  the  addition  of  salt  after  the  sulphating 
period,  at  least  not  in  its  present  construction.  However,  this 
furnace  is  well  adapted  for  quite  a  variety  of  ores,  especially  those 
which  do  not  cake  easily  and  contain  a  considerable  quantity  of 
iron  pyrites. 

(6)  The  Ropp  Furnace.  —  It  is  apparent  that  by  dragging  the 
plows  on  the  bottom  of  the  furnace,  as  in  the  O'Harra  furnace, 


MECHANICAL  ROASTING   FURNACES 


75 


the  bottom  and  the  rabbles  or  plows  will  suffer  much  wear,  and  not 
less  the  chain,  especially  when  highly  sulphureted  ore  is  roasted. 
Alfred  von  der  Ropp  has  obviated  these  obstacles  by  a  very 
ingenious  construction  of  his  furnace.  Figs.  19  and  21  B  repre- 
sent a  horizontal  section,  and  Fig.  21  A  an  elevation  of  his  fur- 
nace, while  Fig.  20  is  a  cross-section  through  fireplace  and  hearth. 


FIG.  19.— HORIZONTAL  SECTION  OF  ROPP  FURNACE. 

It  is  a  one-story  straight  hearth  furnace,  105  ft.  Jong  and  11  ft. 
wide  in  the  clear.  The  hearth  is  longitudinally  divided  into  two 
even  parts  by  a  slot,  which  extends  the  whole  length.  The  two 
sides  of  the  slot  project  above  the  hearth  surface  to  prevent  the 
ore  from  falling  into  it.  The  slot  communicates  with  a  tunnel,  E, 
underneath  the  furnace.  This  tunnel  is  provided  with  a  track 


FIG.  20.  —  CROSS-SECTION  OF  ROPP  FURNACE. 

on  which  travels  a  four-wheeled  truck  or  carriage.  To  this  truck  is 
fastened  an  iron  arm  which  extends  through  the  slot  into  the  fur- 
nace, and  to  the  end  of  which  is  attached  a  cross  arm  extending 
over  the  whole  width  of  the  hearth.  This  cross  arm  is  provided 
with  adjustable  rabbles,  which  are  designed  not  only  to  stir  the 
ore,  but  also  to  move  it  gradually  forward.  An  endless  wire 


76 


HYDROMETALLURGY   OF  SILVER 


MECHANICAL   ROASTING   FURNACES  77 

rope  moves  the  truck  and  with  it  the  rabble  arrangement.  The 
truck,  after  passing  underneath  the  furnace,  turns  a  curve  and 
returns  on  the  outside  to  the  other  end  of  the  furnace,  where  it 
enters  the  furnace  again.  On  this  track  run  four  carriages  or 
trucks,  which  are  placed  at  even  distances  from  each  other. 
They  are  fastened  to  the  wire  rope.  The  rabbles  are  so  arranged 
that  if  the  preceding  rake  turns  the  ore  toward  the  right  the 
following  one  will  turn  it  to  the  left.  The  rabbles  can  be  set 
more  or  less  slanting  as  circumstances  require,  and  also  can  be 
lowered  or  raised.  The  tunnel  underneath  the  furnace  is  high 
enough  for  a  man  to  enter  and  to  pass  underneath  the  truck  in 
case  any  repairs  should  be  required.  Three  fireplaces  provide 
the  necessary  heat.  They  are  all  on  one  side  of  the  furnace  and 
are  constructed  as  shown  in  Fig.  20.  The  feeding  of  the  ore  is 
done  automatically  by  Challenge  feeders,  at  one  end  of  the  fur- 
nace. At  the  opposite  end  the  roasted  ore  is  brought  out,  each 
rake  pushing  forward  a  certain  amount  and  dumping  it  into  two 
iron  cars.  The  two  ends  of  the  furnace  are  closed,  like  the 
O'Harra  furnace,  by  swinging  doors,  which  are  opened  by 
fenders  attached  to  the  rake  and  which  drop  back  in  position 
after  the  rake  has  passed.  Numerous  doors  on  both  sides  of 
the  furnace  serve  for  regulating  the  admission  of  air  and  give 
access  for  cleaning  the  hearth. 

The  driving  mechanism  is  of  simple  construction.  Two  bevel 
gears  transmit  the  power  to  a  horizontal  sheave  around  which 
the  steel  wire  rope  travels.  A  similar  sheave  is  arranged  on  the 
opposite  end  of  the  furnace.  The  rope  is  entirely  outside  the 
furnace  and  is  therefore  perfectly  protected  from  the  heat,  and 
so  are  the  carriages.  The  rakes  returning  on  the  outside  of  the 
furnace  for  so  long  a  distance  have  sufficiently  cooled  at  the  time 
when  they  enter  the  furnace  again. 

It  is  claimed  that  the  furnace  has  a  capacity  of  36  tons  in 
twenty-four  hours  roasting  an  ore  which  contains  20  per  cent, 
sulphur,  8  per  cent,  lead  and  17|  per  cent.  zinc.  This  furnace  is 
also  excellently  adapted  for  oxidizing  roasting  of  iron  pyrites 
and  copper  matte. 

With  respect  to  the  roasting  conditions  of  this  furnace  they 
are  the  same  as  prevail  in  the  O'Harra,  but  the  mechanical  con- 
struction is  much  superior  to  that  furnace. 

(c)  The    Howell-White    Furnace,  —  About    the   simplest    and 


78 


HYDROMETALLURGY  OF  SILVER 


cheapest  continuous  mechanical  furnace,  and  at  the  same  time 
perhaps  the  most  effective,  is  the  •  Howell-White.  It  is  very 
simple  in  construction,  requires  little  repair  and  very  little  power 
to  run  it,  while  its  capacity  is  quite  large. 

This  furnace  is  a  revolving  cylinder,  open  on  both  ends.  One 
of  them  is  connected  with  the  fireplace  while  the  other  enters 
into  the  flue.  The  cylinder  is  made  of  cast  iron  in  flanged  sec- 
tions, which  in  setting  up  are  bolted  together.  The  two  sections 
nearest  to  the  fire  are  lined  with  bricks  to  protect  the  iron,  and 
in  order  to  have  the  inside  diameter,  after  lining,  the  same  through 
the  whole  length  of  the  furnace,  the  diameter  of  these  two  sec- 
tions is  10  inches  larger  than  the  balance.  The  furnace  rests  on 
the  top  of  five  wheels  all  in  one  line  and  properly  divided,  as 
can  be  seen  in  Fig.  22.  It  is  kept  in  place  by  four  rollers  which 


FIG.  22.  —  HOWELL-WHITE  FURNACE. 

touch  the  furnace  half-way  up  its  diameter,  which  are  kept  there 
by  strong,  stiff  iron  frames.  The  furnace  is  made  to  revolve  on 
top  of  the  wheels  to  diminish  friction  and  consequently  to  make 
it  turn  more  easily.  It  has  a  slight  inclination  toward  the  fire. 
The  flue  end  of  the  cylinder  is  provided  with  a  flange  to  prevent 
the  ore  from  falling  into  the  dust-chamber.  The  feeding  is  done 
through  an  inclined  cast-iron  spout,  which  passes  through  the 
arch  of  the  dust-chamber  and  extends  into  the  furnace  to  within 
five  or  six  inches  of  the  periphery.  The  feeding  is  done  from  a 
hopper  by  means  of  a  worm,  the  speed  of  which  is  adjustable. 
To  the  inside  of  the  unlined  part  of  the  cylinder  are  riveted 
several  iron  ribs  in  the  shape  of  a  spiral.  The  object  of  these  ribs 
is  to  lift  the  ore,  when  the  furnace  is  revolving,  and  to  shower  it 
through  the  flame.  As  this  causes  much  dust,  which  is  carried 
away  by  the  draft,  an  auxiliary  fireplace  is  arranged  at  that  end, 


MECHANICAL  ROASTING  FURNACES 


79 


the  flame  entering  the  dust-chamber  right  under  the  end  of  the 
furnace. 

The  roasted  ore  leaves  the  furnace  gradually  at  the  fire  end 
and  drops  into  a  vault  underground  but  conveniently  accessible 
from  the  cooling  floor.  For  this  purpose  a  space  about  12  in. 
wide  is  left  between  the  fire-bridge  and  the  furnace  end.  The 


FIG.  23.— HOWELL  FURNACE,  DISCHARGE  END  AND  ORE-VAULT. 

revolving  speed  of  the  furnace  is  adjustable;  in  most  cases  2J  to  3 
revolutions  per  minute  is  sufficient.  The  forward  movement  of 
the  ore  in  the  cylinder  is  caused  by  the  revolving  motion,  and 
the  speed,  therefore,  regulates  the  amount  of  ore  which  passes 
through  and  the  amount  which  is  retained  in  the  cylinder  while 
in  operation.  Fig.  23  represents  a  longitudinal  section  showing 


80  HYDROMETALLURGY  OF  SILVER 

the  relative  position  of  fireplace  and  furnace,  the  ore-drop  and 
the  ore- vault. 

The  constant  showering  of  the  ore,  brought  up  by  the  ribs, 
through  the  draft  causes  a  separation  of  dust  and  sand,  the  dust 
being  carried  into  the  dust-chamber.  The  amount  of  dust  de- 
posited in  the  chambers  as  compared  with  that  of  the  roasted  ore 
dropping  into  the  vault  is  from  30  to  50  per  cent.  The  dust  is 
well  roasted,  provided  proper  attention  is  paid  to  the  auxiliary 
fire  and  that  it  is  kept  strong  enough.  This,  however,  can  only 
be  done  if  the  fire  is  intrusted  to  an  extra  man.  If  it  is  made 
part  of  the  duty  of  the  man  attending  the  fire  at  the  discharge 
end,  the  auxiliary  will  always  be  more  or  less  neglected.  One 
man  at  each  end  can  attend  to  three  furnaces. 

There  are  ores,  however,  which  will  not  stand  entering  the 
sudden  heat  of  the  auxiliary  without  caking.  Even  if  the  dust 
is  roasted  well  by  the  auxiliary  fire,  the  formation  of  so  much  of 
it  is  very  annoying  and  inconvenient.  Actually  there  is  no  need 
to  make  so  much  dust  with  this  furnace. 

This  excessive  amount  of  dust  is  caused  by  the  ribs,  which 
produce  a  shower  of  ore  through  the  swiftly  moving  gases.  I 
found  that  these  ribs  are  not  necessary.  There  is  no  accumulation 
of  ore  in  this  furnace;  in  fact  there  is  only  a  comparatively  small 
stream  of  ore  passing  through,  even  if  roasting  is  conducted  at 
the  rate  of  30  tons  in  twenty-four  hours.  By  the  revolving 
motion  of  the  furnace,  this  thin  layer  of  ore  is  made  to  expose 
continually  new  particles  to  the  action  of  heat  and  air,  and  the 
ore  at  the  discharge  end  will  be  found  just  as  well  roasted  with- 
out as  with  ribs;  in  fact  better,  because  it  will  contain  a  larger 
percentage  of  fine  material.  Besides,  the  furnace  after  being 
incrusted  offers  a  rough  surface  to  the  ore,  which  prevents  its  sli- 
ding swiftly  and  spreads  it  over  quite  a  large  surface.  By  remov- 
ing the  spiral  ribs  from  the  Howell  furnaces  of  the  Cusihuiriachic 
Mining  Company,  Mexico,  I  very  much  diminished  the  formation 
of  dust ;  in  fact  so  much  so  that  the  maintenance  of  the  auxiliary 
fire  was  not  justified,  and  was  abandoned  altogether.  The  dust 
from  right  behind  the  furnace  was  removed  twice  a  day  and  ele- 
vated to  the  feed-bin,  thus  mixing  it  with  the  crude  ore  and 
feeding  it  into  the  furnace  again. 

To  have  an  elevator  between  each  two  furnaces,  which  can 
be  made  to  discharge  into  either  of  the  two  feed-hoppers,  is  very 


MECHANICAL  ROASTING  FURNACES  81 

convenient,  not  only  for  elevating  the  dust  from  the  chambers, 
but  also  to  elevate  the  ore-sweepings,  which  always  accumulate 
around  a  roasting  furnace. 

This  furnace  radiates  considerable  heat,  which  affects  the 
driving-belt  and  shortens  its  life;  it  is,  therefore,  much  better 
to  drive  the  furnace  by  means  of  sprockets  and  link  chain. 

It  ought  to  be  mentioned  that  there  are  two  vaults  beneath 
each  furnace  which  can  be  filled  alternately  by  turning  a  wing 
which  is  placed  right  under  the  roasted  ore  drop.  On  each  side 
of  this  drop  there  should  be  a  small  door  to  permit  the  entrance 
of  air  into  the  furnace  and  to  give  access  for  tools  in  case  it  is 
necessary.  » 

If  one  of  the  vaults  is  filled  with  ore,  it  is  advisable  not  to 
discharge  it  until  the  other  is  nearly  filled.  This  gives  the  ore 
an  opportunity  to  improve  in  chlorination. 

When  the  crust  in  the  furnace  becomes  too  thick  it  can  be 
easily  removed  by  inserting  a  number  of  bricks  at  the  flue  end. 
In  revolving  the  bricks  will  shave  off  the  soft  crust  and  bring 
it  out.  When  the  bricks  are  charged  the  feed  has  to  be  stopped 
for  a  while;  but  as  soon  as  they  have  moved  away  4  or  5  ft. 
from  the  end,  charging  can  be  commenced  again. 

(d)  0.  Hofmann's  Modified  Howell  Furnace.  —  The  Howell 
furnace  is  a  very  efficient  furnace;  its  first  cost  is  small  as  com- 
pared to  its  roasting  capacity;  it  does  not  need  many  repairs,  and 
requires  but  very  little  manual  labor.  The  furnace  does  not 
carry  much  ore  at  a  time,  and  is  therefore  not  excessively  heavy; 
the  way  it  is  arranged,  revolving  on  top  of  wheels,  the  friction  is 
much  reduced,  so  that  the  furnace  revolves  very  easily  and  not 
more  than  two,  perhaps  two  and  a  half,  horse-power  is  required. 
All  these  favorable  qualities  make  the  Howell  furnace  a  very 
desirable  one;  but  it  has  the  drawback,  in  common  with  all  the 
continuous  furnaces,  that  the  salt  has  to  be  mixed  with  the  ore 
before  it  enters  the  furnace.  We  have  seen  above  that  there 
are  ores  which  cannot  be  chloridized  unless  the  salt  is  added 
later  during  roasting,  and  as  the  construction  of  the  Howell 
furnace  does  not  permit  of  this,  it  makes  it  unfit  for  such  ores, 
which  is  to  be  regretted  considering  the  numerous  advantagous 
features  of  this  furnace. 

In  my  metallurgical  investigations  of  the  heavy  argentiferous 
zinc-lead  ores  of  the  San  Francisco  del  Oro  mine,  Chihuahua, 


82  HYDROMETALLURGY  OF  SILVER 

Mexico,  I  was  confronted,  besides  other  roasting  questions,  with 
the  problem  as  to  the  best  method  of  chloridizing  these  ores  in  a 
Howell  furnace.  By  mixing  the  salt  with  the  ore  in  the  stamp 
battery  the  ore  became  sticky,  incrusted  the  furnace  rapidly,  and 
when  it  left  the  furnace  consisted  mostly  of  lumps  without  being 
much  chloridized.  If  the  ore  was  charged  without  salt,  it  re- 
mained loose  and  sandy,  but  it  dusted  so  much  that  by  the  draft 
an  almost  perfect  separation  of  coarse  and  fine  took  place,  the 
latter  being  carried  into  the  dust-chamber,  while  the  coarse  sand 
dropped  in  the  pit,  but,  on  account  of  its  large  percentage  of 
lead  and  zinc  blende,  insufficiently  oxidized,  so  that  the  salt, 
which  was  added  from  time  to  time  to  the  ore  in  the  pit  and 
stirred,  had  but  very  little  chloridizing  effect.  The  best  chlorina- 
tion  obtained  was  only  29  per  cent.  To  diminish  this  separation 
a  small  percentage  of  salt,  from  1  to  2  per  cent,  was  added  in  the 
battery.  The  effect  was  remarkable;  without  balling  or  incrusting 
the  furnace  the  dusting  was  practically  stopped.  When  the 
balance  of  the  salt  was  added  in  the  pit  a  chlorination  of  67  per 
cent,  was  obtained. 

Based  on  this  observation  a  modification  of  the  arrangement 
in  front  of  the  furnace  was  made.  Between  the  fireplace  and 
the  discharge  end  of  the  furnace  a  shallow  pit  was  inserted,  which 
was  in  communication  with  a  reverberatory  furnace  6  by  8  ft. 
The  bottoms  of  both  were  on  the  same  level.  The  fireplace  of 
the  reverberatory  was  only  24  in.  wide.  The  gases  from  the 
reverberatory  passed  through  the  pit  and  the  furnace.  When 
a  charge  of  about  1400  Ib.  had  accumulated  in  the  pit  the  same 
was  pushed  into  the  reverberatory,  salt  added  and  well  mixed. 
There  the  ore  was  kept  until  another  charge  had  accumulated  in 
the  pit.  By  this  modification  very  satisfactory  chlorination  was 
obtained.  A  very  low  fire  was  kept  on  both  the  fireplaces. 

(e)  The  Stetefeldt  Furnace.  —  This  furnace  consists  of  an 
upright  shaft  30  to  40  ft.  high,  which  is  connected  near  the  top 
with  a  descending  flue.  On  top  of  the  shaft  is  a  feeding  machine, 
which  showers  the  ore  into  the  shaft.  In  the  lower  part  are  two 
fireplaces  opposite  each  other,  so  that  the  descending  ore  meets 
the  hot  ascending  gases.  The  principle  on  which  the  con- 
struction of  this  furnace  is  based  is  at  variance  with  that  of  any 
other  chloridizing  furnace.  The  ore  falls  very  finely  divided 
through  a  glowing  atmosphere  of  chlorine,  sulphurous  acid, 


MECHANICAL  ROASTING  FURNACES  83 

oxygen  and  fire  gases,  metal  chlorides  and  volatilized  salt,  and 
the  whole  complicated  reactions  which  take  place  in  chloridizing 
roasting  are  completed  in  the  incredibly  short  time  of  a  very  few 
seconds.  Mr.  Stetefeldt  derived  his  idea  from  the  Gerstenhofer 
pyrites  roaster,  in  which  the  ore  is  fed  into  a  shaft,  wherein 
numerous  shelves  are  so  arranged  that  the  ore  drops  from  one 
shelf  to  the  other,  resting  on  each  for  some  time.  The  experi- 
mental furnace  Mr.  Stetefeldt  built  was  of  such  a  construction 
arranged  so  that  fire  gases  were  permitted  to  enter  the  shaft.  The 
shelves,  however,  caused  much  inconvenience  by  incrustation, 
etc.,  and  by  observations  he  made  during  these  experiments  he 
thought  it  justifiable  to  repeat  the  experiment  without  the  use 
of  any  shelves.  The  results  were  so  gratifying  that  he,  adhering 
to  this  new  principle,  gradually  developed  his  furnace  to  the 
present  much  improved  construction. 

Fig.  24  represents  a  vertical  section  of  this  furnace.  A,  shaft; 
G,  returning  flue;  K,  receiver  or  hopper  forming  the  bottom  of 
the  shaft;  C,  C,  the  two  fireplaces;  T,  slit  for  the  fire  gases;  U,  air 
ducts  to  produce  a  perfect  combustion,  and  at  the  same  time  to 
cool  the  fire  arch;  E,  ash-pit,  the  iron  door  of  which  is  provided 
with  a  slide  to  regulate  the  air  inlet ;  0,  0,  doors  for  the  insertion 
of  tools  for  cleaning  the  walls.  The  returning  flue,  G,  is  provided 
with  doors,  R,  which  serve  to  clean  it.  D  is  an  auxiliary  fire- 
place, which  is  constructed  like  the  fireplaces  of  the  shaft  and 
which  serves  to  roast  the  large  amount  of  flue-dust  which  this 
furnace  makes.  Passing  the  chamber  H  the  dust  enters  V  and 
the  larger  part  of  it  settles  in  the  bottom  hoppers,  /,  /,  from 
which  it  is  drawn  into  iron  cars  by  moving  the  dampers  S,  S. 
The  rest  of  the  dust  is  collected  in  a  number  of  dust-chambers,  Q, 
which  are  connected  with  the  chimney  by  means  of  a  long  flue. 
P,  P  are  doors  for  observation  and  cleaning.  Below  the  shaft- 
hopper,  K,  is  the  slide  door  L.  By  pulling  this  slide  the  accumu- 
lated roasted  ore  drops  into  a  large  iron  car.  B  is  a  cast-iron 
frame  with  water-jacket  on  top  of  the  shaft  on  which  the  ore- 
feeding  machine  is  placed.  Above  it  is  an  ore  bin  (not  shown 
in  the  diagram)  from  which  the  ore  is  fed  into  the  machine  by 
means  of  a  worm. 

The  feeding  machine  is  shown  in  Fig.  25.  A  is  a  cast-iron, 
water-cooled  frame  placed  on  top  of  the  shaft  and  provided  with 
the  damper,  B,  which  is  withdrawn  when  the  furnace  is  in  opera- 


84 


HYDROMETALLURGY  OF  SILVER 


tion,  but  which  is  inserted  if  the  feeding  machine  stops  for  any 
repair  or  for  exchange  of  screens.  C  is  a  cast-iron  grate  to  which 
on  the  upper  side  is  fastened  the  punched  screen,  D,  which  is 
made  of  a  steel  plate  with  holes  of  an  eighth  of  an  inch.  Above 
the  punched  screen  is  placed  a  frame,  E,  to  the  bottom  of  which 
is  fastened  a  coarse  wire  screen,  F,  with  about  three  meshes  to 


FIG.  24.— STETEFELDT  FURNACE. 

the  inch,  made  of  heavy  wire.  The  frame,  E,  rests  upon  friction 
rollers,  G.  The  brackets,  H,  which  hold  the  friction  rollers,  can 
be  raised  or  lowered  by  set-screws  so  that  the  wire  screen  can  be 
brought  closer  or  less  close  to  the  punched  screen.  The  brackets 
K  carry  an  eccentric  shaft,  by  which  an  oscillating  motion  is 
given  to  the  frame  E.  To  the  brackets  N  are  fastened  transverse 
stationary  blades,  0,  which  extend  close  to  the  wire  screen  and 


MECHANICAL  ROASTING  FURNACES 


85 


which  can  be  adjusted  by  the  nuts  P.  These  blades  keep  the  ore 
uniformly  spread  over  the  screen  when  the  machine  is  in  motion. 
The  ore  is  usually  crushed  through  a  40-mesh  screen. 

Before  starting  the  feed  the  furnace  has  to  be  well  heated, 
which  takes  thirty-six  to  forty  hours.  After  the  speed  of  the 
feeder  is  set  and  the  proper  temperature  is  ascertained  by  obser- 
vation and  numerous  assays,  the  roasting  itself  requires  but  very 
little  attention  beyond  the  maintenance  of  a  uniform  temperature. 


FIG.  25.  — FEEDING  MACHINE,  STETEFELDT  FURNACE. 

The  capacity  of  the  furnace  is  very  large,  roasting,  according 
to  the  size  of  it  and  the  character  of  the  ore,  from  20  to  50  tons 
in  twenty-four  hours.  The  consumption  of  fuel  is  rather  small. 
With  one  cord  of  wood  about  8  tons  of  ore  can  be  roasted.  The 
loss  of  silver  by  volatilization  is  less  than  if  the  same  ore  was 
roasted  at  the  same  heat  in  another  furnace,  because,  according 
to  Plattner,  the  loss  of  silver  does  not  depend  merely  on  the 
temperature  the  ore  is  roasted  at  and  the  character  of  the  ore, 
but  also  on  the  length  of  time  the  ore  is  subjected  to  the  roasting 
temperature.  In  the  Stetefeldt  furnace  the  ore  is  roasted  almost 
instantaneously  and  does  not  suffer  that  part  of  the  loss  which 
is  caused  by  long  heating. 

This  furnace  is  well  adapted  to  roast  ores  not  heavily  charged 
with  sulphides,  containing  5  to  8  per  cent,  sulphur,  or  even  10 
per  cent.,  if  no,  or  only  a  small  percentage  of,  galena  and  zinc 


86  HYDROMETALLURGY  OF  SILVER 

blende  is  present.  Ores  heavily  charged  with  sulphides  are  not 
suitable,  especially  if  they  contain  a  large  percentage  of  galena 
and  zinc  blende.  These  minerals  need  a  low,  gradually  increas- 
ing temperature,  which  conditions  cannot  be  maintained  in  the 
Stetefeldt  furnace.  If  highly  sulphureted  ore  is  fed  into  the 
shaft  the  temperature  in  the  upper  part  of  the  shaft  extend- 
ing closely  to  the  feeding  machine  becomes  intense,  and  the 
conditions  required  for  roasting  such  ores  become  reversed. 
Instead  of  being  exposed  to  a  low  and  then  gradually  increasing 
temperature,  the  ore  encounters  the  hottest  zone  first,  and  com- 
bustion is  so  rapid  that  when  it  drops  to  the  bottom  it  will  be  found 
to  consist  mostly  of  minute  globules,  formed  by  partial  melting 
of  the  ore  particles.  Besides,  much  crust  is  formed  at  the  bottom 
of  the  shaft  as  well  as  on  the  sides,  which  keeps  falling  down  in 
large  chunks.  Even  the  iron  grate  on  which  the  punched  screen 
of  the  feeding  machine  rests  becomes  clogged  with  incrustation, 
which  occurs  so  frequently  as  to  make  a  regular  feeding  impos- 
sible. This  was  the  reason  that  while  a  large  number  of  this 
type  of  furnace  were  in  successful  operation  on  the  Pacific  slope, 
in  Nevada,  California,  Utah,  etc.,  all  the  numerous  attempts  to 
work  the  more  mineralized  ores  of  Mexico  were  failures. 

For  the  proper  ore,  this  furnace  is  undoubtedly  the  cheapest 
to  roast  ores,  especially  for  those  which,  on  account  of  their 
small  percentage  of  sulphur,  cause  a  large  consumption  of  fiiel. 
The  conditions  necessary  to  heat  such  ores  quickly  and  effectively 
are  excellent,  because  every  particle  is  exposed  to  the  heat  when 
it  passes  in  a  shower  through  the  flame. 

G.  Kiistel  had  once  to  roast  in  a  Stetefeldt  furnace  ores  which 
were  too  poor  in  sulphur  to  produce  sufficient  chlorine.  Sul- 
phureted ore  to  mix  with  the  dry  ore  could  not  be  procured, 
either.  Mr.  Kiistel  solved  the  problem  by  burning  brimstone  in 
a  cast-iron  pan  and  conveying  the  gas  into  the  furnace  near  the 
bottom  of  the  shaft.  This  could  not  have  been  done  successfully 
with  any  other  furnace. 


IX 

COLLECTING  THE  FLUE-DUST 

THE  formation  of  dust  in  roasting  is  unavoidable.  In  some 
of  the  furnaces  the  amount  of  dust  carried  off  by  the  draft  is 
comparatively  small,  as  in  the  reverberatory  and  Bruckner, 
while  in  others,  like  the  White-Howell  and  the  Stetefeldt,  it  is 
excessive.  The  dust  consists  of  ore  particles,  more  or  less  roasted, 
carried  out  of  the  furnace  by  the  draft,  and  of  condensed  fumes 
of  volatilized  metal  chlorides  and  oxides.  The  former  are  much 
easier  collected  than  the  latter. 

The  problem  of  collecting  this  dust  is  quite  a  difficult  one,  but 
is  very  interesting  and  of  great  importance  in  metallurgy.  The 
idea  of  collecting  the  dust  by  conveying  the  gases  into  large 
chambers  in  order  to  reduce  the  speed  of  their  movement,  thus 
giving  time  for  the  dust  to  settle,  is  not  a  correct  one.  Close 
observations  have  shown  that  in  such  a  chamber  a  much  larger 
accumulation  of  dust  will  be  found  close  to  the  walls  than  on 
other  parts  of  the  floor.  Besides,  the  walls  will  be  found  to  be 
covered  with  scale-like  formations  of  the  dust,  the  pointed 
part  of  these  scales  turned  against  the  current  of  the  gases.  When 
these  scales  become  too  heavy  they  peel  off  and  drop  to  the  floor, 
hence  the  larger  accumulation  of  dust  on  the  floor  near  the  foot 
of  the  walls.  The  cause  of  this  greater  precipitation  can  be  no 
other  than  the  friction  between  the  dust-charged  gases  and  the 
walls.  It  can  also  be  observed  that  wherever  the  flue  makes 
a  sharp  turn  the  precipitation  of  dust  is  greater,  because  the 
friction  between  wall  and  gases  is  much  greater  if  the  latter  are 
forced  to  make  a  sudden  change  in  their  course  than  if  they  are 
permitted  to  follow  a  straight-line  course.  Furthermore,  it  can 
be  observed  that,  if  an  obstacle  is  placed  in  the  flue,  against 
which  the  moving  gases  have  to  strike,  it  will  cause  a  precipita- 
tion of  dust  which  increases  with  the  swiftness  of  the  gas  current. 

87 


88 


HYDROMETALLURGY  OF  SILVER 


These  observations  demonstrate  that  a  slow  movement  of 
the  gases,  as  attained  with  large  dust-chambers,  is  by  far  not 
so  effective  as  a  swift  movement  with  increased  friction,  and 
therefore  much  better  results  will  be  obtained  if  the  dust-chambers 
are  so  constructed  as  to  offer  a  large  wall  surface  as  compared 
with  the  area  of  the  cross-section  of  the  chambers,  and  if  the 
same  are  arranged  in  a  zigzag  fashion. 

Based  on  the  above  observations  I  have  devised  and  con- 
structed a  dust-collecting  arrangement  which  is  very  effective  and 
gives  much  satisfaction. 

14-8fc- >j 


L- 


FIG.  26.— VERTICAL  SECTION  OF  HOFMANN  DUST  COLLECTOR, 

0.  Hofmanri's  Flue-Dust  Collector.  —  Fig.  26  represents  in  a  ver- 
tical section  the  dust-collecting  arrangement  in  connection  with 
a  White-Howell  roasting  furnace.  A,  feed  end  of  the  furnace; 
J5,  first  chamber,  the  end  of  the  furnace  projecting  a  few  inches 
into  it.  Here  the  coarsest  part  of  the  dust  accumulates  and  is 
removed  from  time  to  time  through  the  door  H.  From  here  the 
gases  pass  through  the  opening,  K,  into  the  second  chamber,  C, 
and  from  there  through  the  arch,  M,  into  the  collecting  shaft,  D, 
which  they  ascend,  leaving  it  through  the  flue,  F,  and  entering 


COLLECTING  THE   FLUE-DUST 


89 


the  second  collecting  shaft,  E,  in  which  they  descend,  leaving 
the  same  through  G,  which  makes  connection  with  the  general 
flue  leading  to  the  chimney. 


FIG.  27. —DETAILS  OF  BARS  AND  BEARINGS,  HOFMANN 
DUST  COLLECTOR. 

In  the  collecting  shafts,  cast-iron  double  channel  irons  are 
arranged  in  rows,  leaving  a  space  3  in.  wide  between  each  two. 
These  channels  are  closed  at  each  end  (Fig.  27)  and  extend  as 
round  bars  If  in.  in  diameter  and  1  ft.  8|  in.  long  at  the  front 
and  3  in.  long  at  the  opposite  side.  These  two  cylindrical  exten- 


4  — , 


Section,  on  Line  A?B 


T 

1  /-:-;•-  1 

?~T 

'  ' 

f    1* 

I  v-y  i 

i 

[ft 

Scale  l^*= 

Bearing  for  Bar         Scale  1J$  - 1  ft. 

FIG.  28.  — DETAILS  OF  BARS  AND  BEARINGS, 
HOFMANN  DUST  COLLECTOR. 

sions  rest  in  cast-iron  bearings,  as  shown  in  Fig.  28.  The  longer 
bar  extends  through  the  front  wall  of  the  shaft,  and  is  cast  square 
at  the  very  end,  which  part  projects  out  of  the  wall,  while  the 
shorter  rests  in  a  bearing  inserted  in  the  back  wall  of  the  shaft. 
These  channels  can  be  turned  by  a  socket  wrench  slipped  over 


90 


HYDROMETALLURGY  OF  SILVER 


the  square  end.  Each  row  is  10  in.  above  the  other,  and  the 
position  of  the  channels  is  such  that  the  channels  of  one  row  are 
placed  right  above  the  open  spaces  of  the  row  below,  as  shown  in 
Fig.  29.  The  bearings  are  1J  in.  in  diameter  while  the  cylindri- 
cal parts  of  the  bars  are  only  If  in.,  in  order  to  allow  room 
for  expansion  and  to  permit  an  easy  turning  of  the  bars  when 
hot. 

The  gases  from  the  furnace  entering  the  shaft  D,  through  M 
(Fig.  26),  in  ascending  partly  strike  the  channels  of  the  first  row 
and  partly  pass  through  the  open  spaces.  That  part,  however, 
which  strikes  the  channel  rebounds  and  is  also  forced  to  pass 
through  the  open  spaces.  Passing  through,  they  strike  against 


M- 8*- 


i^.  29.  — POSITION  OF  BARS,  HOFMANN  DUST  COLLECTOR. 

Bearing  should  be  made  If  in.  diameter,  or  J  in.  larger  than  the  bar,  so  that  the 

latter  can  be  easily  turned  when  hot. 

the  channels  of  the  next  row  above,  rebound,  and  force  their  way 
again  through  the  open  spaces.  This  is  repeated  until  the  gases 
leave  the  top  row  and  enter  the  second  shaft,  E.  Here  the  same 
play  of  the  gases  takes  place,  only  that  they  have  to  descend 
through  the  shaft. 

It  is  apparent  that  the  numerous  objects  placed  in  the  path 
of  the  gases  and  against  which  they  have  to  strike,  and  the 
large  surface  which  they  offer  for  friction  will  produce  a  very 
effective  precipitation  of  the  fumes  and  dust.  This  is  actually 
the  case.  I  erected  at  the  works  of  the  United  Zinc  and  Chemical 
Company,  Argentine,  Kansas,  a  system  of  three  such  collecting 
shafts.  Blende-pyrite  ore  was  roasted  in  mechanical  furnaces  to 


COLLECTING  THE   FLUE-DUST  91 

produce  sulphur  dioxide  gas  for  the  manufacture  of  sulphuric 
acid,  and  it  was  of  importance  that  the  gases  should  enter  the 
Gay-Lussac  tower  as  free  of  dust  as  possible.  After  the  system 
was  in  operation  for  some  time  an  investigation  was  made  as  to 
its  efficiency.  The  fumes  passed  first  through  a  down  flue  and  a 
piece  of  straight  flue,  in  which  the  coarse  dust  accumulated  before 
it  entered  the  first  shaft.  It  was  found  that  the  dust  on  the 
bottom  of  the  first  shaft  had  the  color  of  the  roasted  ore,  showing 
that  the  main  part  of  it  consisted  principally  of  mechanical  ore- 
dust.  The  dust  on  the  bottom  of  the  second  shaft  was  of  a 
light  pink  color,  showing  that  the  main  part  of  it  consisted  of 
precipitated  white  fumes  of  lead,  iron  and  zinc.  On  the  bottom 
of  the  third  shaft  the  dust  was  perfectly  white,  showing  that  all 
the  ore-dust  was  precipitated  before  it  .reached  the  third  shaft, 
which  was  proved  by  actual  analysis  of  this  dust,  showing  it  to 
consist  of  the  sulphates  of  lead,  zinc  and  principally  iron,  while 
almost  no  insolubles  were  present.  In  the  first  shaft  were  found 
6.65  cu.  ft.  of  dust;  in  the  second,  4.15  cu.  ft.,-  and  in  the  third 
only  1.55  cu.  ft.  This  very  rapidly  decreasing  volume  of  dust 
found  in  each  succeeding  shaft  illustrates  the  great  efficiency  of 
this  system.  I  may  mention  as  a  further  illustration  that,  at 
the  bottom  of  the  down  flue,  where  the  coarse  settled,  the  layer 
of  dust  was  7  in.  deep.  In  the  piece  of  straight  flue  next  to  the 
first  shaft  the  layer  was  only  1J  in.  thick,  while  in  the  first  shaft 
it  was  8  in. 

The  channels  have  to  be  shaken  and  turned  from  time  to  time, 
depending  on  the  dusting  qualities  of  the  ore  and  furnace.  Once 
or  twice  a  week  will  be  found  sufficient.  The  shaking  and  turning 
should,  of  course,  be  commenced  at  the  top  row.  Fig.  30  shows 
the  dust-collecting  and  flue  arrangement  in  connection  with  two 
White-Howell  furnaces. 

This  dust-collecting  arrangement  is  compact,  and  very  effect- 
ive, and  ought  to  be  inserted  in  all  the  works  where  ores  are 
roasted,  either  in  order  to  regain  the  valuable  dust  and  the  vola- 
tilized silver,  copper,  lead,  etc.,  or  to  prevent  the  entering  of  the 
dust  and  volatilized  substances  into  the  subsequent  chemical 
process. 

A  very  effective  method  of  collecting  the  dust  is  the  bag 
system,  in  which  the  gases  are  forced  by  fans  through  long  bags 
made  of  burlap,  flannel  or  muslin,  which  act  as  filters.  The  gases, 


92 


HYDROMETALLURGY  OF  SILVER 


r1 


COLLECTING  THE  FLUE-DUST  93 

however,  have  to  be  first  cooled  sufficiently  so  as  not  to  ignite 
the  bags,  which  cannot  always  be  easily  and  cheaply  accom- 
plished. Besides,  if,  as  in  a  sulphuric  acid  factory,  the  furnace 
gases  have  to  enter  the  process  hot,  the  bag  system  cannot  be 
applied.  This  system  is  often  used  in  smelting  works. 


.    x 

SULPHATING  ROASTING 

THIS  mode  of  roasting,  which  has  the  object  of  converting 
the  silver  into  silver  sulphate,  in  which  state  it  is  soluble  in  water, 
is  only  used  if  silver  is  to  be  extracted  with  hot  water  by 
Ziervogel's  method. 

The  material  to  be  suitable  for  this  roasting  has  to  consist 
principally  of  copper  and  iron  sulphides,  of  which  the  former  has 
to  predominate,  and  has  to  be  free  of,  or  to  contain  only  in  small 
quantities,  the  sulphides  of  lead,  zinc,  arsenic,  and  antimony. 
For  this  reason  it  is  exclusively  used  for  argentiferous  copper 
matte.  In  this  roasting  the  copper  and  iron  have  to  be  converted 
into  oxides,  while  the  silver  has  to  be  changed  into  a  sulphate. 
The  transformation  of  the  silver  into  sulphate  is  done  almost  exclu- 
sively by  the  sulphuric  acid  fumes  which  result  from  the 
decomposition  of  cupric  sulphate  at  a  higher  heat.  Cupric  sul- 
phate and  ferrous  sulphate  are  formed  in  the  first  stage  of  roasting. 
Ferrous  sulphate  is  decomposed  at  a  much  lower  temperature 
than  cupric  sulphate,  in  fact  at  a  temperature  not  high  enough 
for  the  formation  of  silver  sulphate,  so  that  the  sulphuric  acid 
generated  by  the  decomposition  of  the  ferrous  sulphate  is  of 
very  little  avail  for  the  formation  of  silver  sulphate;  it  is,  however, 
of  great  effect  in  the  formation  of  cupric  sulphate,  which  then,  at 
a  higher  heat,  sulphatizes  the  silver. 

A  certain  percentage  of  iron  sulphide  is  therefore  advanta- 
geous for  this  roasting  process,  but  if  the  iron  sulphide  is  in 
excess  the  formation  of  silver  sulphate,  and  with  it  the  extraction, 
suffers. 

At  Mansfeld,  Germany,  where  this  mode  of  roasting  and  the 
subsequent  extraction  of  the  silver  with  hot  water  was  originated 
by  Mr.  Ziervogel,  the  roasting  charge  consisted  of  sulphur  19.32 
per  cent.,  copper  58,  iron  9.18,  lead  2.48,  zinc  4,31,  manganese 

94 


SULPHATING  ROASTING  95 

0.15,  nickel  0.43,  cobalt  0.83,  silver  0.286,  insoluble  1.08  per 
cent.,  and  permitted  an  extraction  of  91  per  cent,  of  the  silver, 
while  at  Schemnitz,  Hungary,  a  matte  containing  88  per  cent, 
iron  sulphide  and  only  1.5  per  cent.,  of  copper  sulphide,  which 
was  tried  by  this  method,  permitted  only  an  extraction  of  73  to 
75  per  cent,  of  the  silver. 

This  roasting  is  a  very  delicate  process  and  has  to  be  con- 
ducted with  great  care  and  skill,  otherwise  inferior  results  will  be 
obtained. 

At  Mansfeld  the  roasting  is  done  in  a  two-story  reverberatory 
furnace.  The  operations  are  as  follows: 

Six  hundred  pounds  of  pulverized  copper  matte  are  charged 
on  the  upper  hearth,  spread,  and  about  5  Ib.  of  slacked  bituminous 
coal  scattered  over  the  charge,  and  stirred.  This  addition  of 
coal  is  made  merely  to  help  heat  the  charge  in  order  to  hasten 
the  operation.  A  matte  richer  in  iron  sulphide  does  not  need 
the  addition  of  coal,  because  iron  sulphide  ignites  quicker  than 
copper  sulphide.  The  charge  is  stirred  continually,  and  the 
lumps  which  form  have  to  be  mashed  with  the  furnace  shovel. 
They  are  soft  and  easily  mashed,  and  are  more  numerous  if  the 
material  contains  more  iron  than  if  it  is  poorer  in  iron.  They 
are  caused  by  the  conversion  of  the  ferrous  sulphate  into  basic 
ferric  sulphate,  which  melts  easily.  In  this  period  the  iron 
oxidizes  before  the  copper,  and  by  the  action  of  the  sulphuric 
acid  changes  into  ferrous  sulphate,  which  later  at  an  increased 
heat  gives  off  sulphuric  acid  fumes  and  changes  into  ferric  oxide 
and  basic  ferric  sulphate.  The  copper  sulphide  is  converted  into 
cupric  sulphate,  but  more  by  the  acid  fumes  of  the  ferrous  salt 
than  by  the  action  of  the  air. 

The  time  of  roasting  on  the  upper  hearth  is  governed  by  the 
time  required  to  finish  the  charge  on  the  lower  hearth.  During 
this  period,  which  lasts  from  five  and  one-half  to  six  hours,  the 
charge  has  to  be  turned  twice  so  that  all  parts  of  it  are  exposed  to 
the  same  heat. 

When  the  lower  hearth  is  clear,  25  Ib.  of  slack  bituminous 
coal  is  spread  over  the  charge,  which  then  is  drawn  to  the  drop- 
hole  in  the  bottom  of  the  hearth,  through  which  it  falls  to  the 
lower  hearth.  When  the  coal  is  mixed  with  the  charge,  burning 
gases  are  emanating  from  the  ore. 

At  the  end  of  the  first  half-hour  the  ore  on  the  lower  hearth 


96  HYDROMETALLURGY   OF  SILVER 

commences  to  glow  brighter  than  it  did  on  the  upper  hearth, 
caused  by  the  higher  temperature  kept  here  and  the  further 
oxidation  of  the  sulphur.  The  thickness  of  the  charge,  which  is 
about  2J  in.,  swells,  on  account  of  the  burning  sulphur,  to  3£  or 
4  in.  In  order  not  to  burn  the  coal  in  the  charge  too  quickly  by 
the  action  of  the  air,  and  to  give  it  better  opportunity  to  act  on 
the  salts  in  the  ore,  the  draft  is  very  much  checked  while  the 
charge  is  raked  continually  and  very  briskly,  in  order  to  avoid 
as  much  as  possible  the  formation  of  lumps.  This  is  done  for  an 
hour,  after  which  time  all  the  coal  is  consumed.  Then  the  charge 
is  turned,  the  part  from  the  hotter  place  to  the  cooler,  and  that 
from  the  cooler  to  the  hotter  place.  After  this  the  draft  is  in- 
creased to  its  full  capacity  in  order  to  produce  a  strong  oxidation 
by  the  inflowing  air.  This  cools  the  charge  after  a  while  until  it 
becomes  almost  dark.  To  judge  the  end  of  this  period,  a  sample 
is  taken  from  the  middle  of  the  hearth,  cooled,  the  fine  separated 
from  the  lumps,  and  by  means  of  a  spatula  a  ridge  is  made  of 
the  fine  in  a  porcelain  saucer.  The  saucer  is  held  slightly  inclined, 
and  some  water,  drop  by  drop,  is  poured  behind  the  ridge.  The 
water  is  first  absorbed  by  the  sample,  but  after  being  saturated 
a  clear  liquor  slowly  flows  out  from  the  other  side  of  the  ridge. 
By  the  color  of  this  liquor  and  its  behavior  toward  salt  the  pro- 
gress of  the  roasting  is  judged.  If  the  roasting  was  conducted 
right,  by  this  time  the  liquor  should  have  a  clear  blue  color  and 
by  the  addition  of  some  salt  should  give  a  light  precipitate  of 
silver  chloride,  which  is  a  sign  that  the  silver  sulphating  has 
commenced.  If  the  liquor  has  a  dirty  greenish  color  it  shows 
that  some  ferrous  sulphate  is  still  undecomposed  and  a  continua- 
tion of  the  oxidizing  period  is  required. 

The  coal,  which  is  added  to  the  charge  and  vigorously  raked 
and  mixed  with  the  latter  while  the  draft  is  much  checked,  acts 
on  the  neutral  sulphates,  which  are  converted  into  basic  sulphates 
while  sulphurous  acid  escapes.  .  After  the  dampers  are  opened 
and  the  full  draft  is  given  to  the  furnace,  the  sulphides  which  may 
still  exist  will  be  completely  roasted  and  all  ferrous  sulphate  will 
disappear,  which  is  necessary  to  be  accomplished  before  the  sul- 
phating of  the  silver  takes  place.  The  cuprous  oxide  oxidizes 
to  cupric,  and  at  the  end  of  this  period  the  material  should  con- 
sist of  the  free  oxides  of  iron  and  copper,  basic  salts  of  iron, 
copper  and  zinc,  and  neutral  sulphates  of  copper,  zinc,  man- 


SULPHATING   ROASTING  97 

ganese,  some  silver  oxide  and  nearly  all  the  balance  of  the  silver 
as  sulphide. 

When  it  has  been  ascertained  by  the  above  test  that  the  roast- 
ing has  advanced  to  the  proper  stage,  the  ore  is  ready  for  the 
sulphating  of  the  silver.  It  will  be  remembered  that  at  the  end 
of  the  previous  period  the  material  had  cooled  down  almost  to 
darkness  by  the  increased  draft.  The  temperature  has  to  be 
increased  again,  but  care  is  to  be  taken  that  this  is  done  with  a 
clean  oxidizing  flame,  and  that  the  latter  does  not  touch  the  ore, 
so  that  none  of  the  cupric  oxide  is  reduced  to  cuprous  oxide  or 
to  metallic  copper,  as  both  of  them  would  precipitate  metallic 
silver  during  the  subsequent  lixiviation,  which  silver  would  re- 
main in  the  residues.  For  this  purpose"  very  dry,  thin  limb-wood 
should  be  used  only.  Pine  is  not  to  be  recommended  on  account 
of  its  pitch,  which  causes  a  smoky  flame.  The  charge  has  to  be 
continually  raked.  After  an  hour  it  becomes  dark  red  hot, 
and  later  increases  to  bright  red.  The  strong  fire  has  to  be  kept 
up  uninterruptedly.  If  after  two  and  a  half  hours  the  material 
near  the  fire-bridge  is  completely  roasted,  the  charge  is  turned 
and  roasting  continued  for  one-half  to  three-quarters  of  an  hour. 
The  material  is  properly  roasted  when  the  solution  emerging  from 
the  ridge  of  a  sample  on  the  saucer  is  only  of  a  faint  but  clear 
blue  color,  showing  that  but  little  cupric  sulphate  is  present, 
while  by  an  addition  of  salt  a  heavy  white  precipitate  of  silver 
chloride  is  formed. 

If  the  heat  is  too  high,  some  of  the  silver  sulphate  will  be 
reduced  to  metallic  silver  and  all  the  cupric  sulphate  will  be  de- 
composed, and  therefore  the  liquor  of  the  test  will  be  colorless. 

By  increasing  the  heat  to  bright  red  during  this  period,  the 
fuming  sulphuric  acid  liberated  from  the  neutral  cupric  sulphate 
does  not  act  so  energetically  on  the  silver  sulphide  as  the  sulphuric 
acid  fumes  resulting  from  the  decomposition  of  the  basic  cupric 
sulphate;  hence  the  addition  of  fine  coal  at  the  beginning  of 
this  period. 

The  time  required  for  roasting  a  charge  is  eleven  to  twelve 
hours,  of  which  five  and  a  half  to  six  are  consumed  on  the  lower, 
and,  naturally,  just  as  many  hours  on  the  upper  hearth.  All  in 
all,  about  four  charges  or  2400  Ib.  of  material  will  be  roasted  by 
each  furnace  during  twenty-four  hours. 

The  loss  of  silver  in  Mansfeld  was  found  to  be  7.06  per  cent. 


98  HYDROMETALLURGY  OF  SILVER 

which  was  caused  partly  mechanically  by  flue-dust,  partly  by 
volatilization  of  silver  oxide,  which,  however,  in  the  cooler 
regions  of  the  dust-chambers  decomposed  into  silver  and  oxygen. 
In  the  roasted  matte  91.74  per  cent,  of  the  silver  was  converted 
into  silver  sulphate  and  was  extractable,  while  1.20  per  cent, 
silver  remained  in  the  residues. 


XI 

CHLORIDIZING  OF  ARGENTIFEROUS  ZINC- 
LEAD  ORE 

IN  this  chapter  the  detail  records  are  given  of  investigations 
of  chloridizing  roasting  of  argentiferous  zinc  blende  and  galena 
ore,  which  I  had  the  opportunity  to  make  on  a  large  working 
scale. 

The  chloridizing  roasting  of  this  class  of  ore  had  not  pre- 
viously been  made  the  subject  of  a  thorough  investigation  on  a 
large  scale,  and  the  record  of  such  experiments  and  investiga- 
tions may  be  of  interest  and  practical  value. 

The  San  Francisco  del  Oro  mine  is  situated  near  Santa 
Barbara  and  Parral,  Chihuahua,  Mexico.  The  ore  contains  on 
an  average  26  to  30  oz.  silver  per  ton,  besides  a  trace  of  gold. 
The  principal  silver-bearing  minerals  are  zinc  blende,  of  which 
the  ore  carries  37  per  cent,  and  more,  and  galena,  of  which  it  con- 
tains from  13  to  19J  per  cent.  The  heavy  solid  occurrence  of 
the  ore  and  the  great  width  of  the  vein  permits  very  cheap 
mining.  The  cost  per  ton  does  not  exceed  $1.50  Mexican  cur- 
rency, including  hoisting  and  a  slight  hand-assorting. 

(a)  Zinc  Blende.  —  The  fine-grained  black  zinc  blende  pre- 
dominates,  containing  about  25  oz.   silver  per  ton;  but  there 
occurs  also  a  brown  blende  of  a  peculiar  luster,  somewhat  resem- 
bling bronze-colored  mica.     It  is  richer  in  silver  than  the  black 
blende,  assaying  from  55  to  70  oz.  silver.     The  blende  contains 
considerable  cadmium. 

(b)  Galena,  with  40  to  50  oz.  silver  per  ton,  is  finely  im- 
pregnated in  the  zinc  blende,  and  can  scarcely  be  detected  with 
the  eye,  and  only  a  very  small  portion  occurs  as  defined  galena, 
which  makes  it  impracticable  to  lessen  the  percentage  of  lead 
in  the  ore  by  hand-sorting.     An  attempt  was  made  to  sort  out 
the  lead  ore  for  shipping,  but  it  was  soon  abandoned  because  the 

99 


100 


HYDROMETALLURGY  OF  SILVER 


amount  of  pure  lead  ore  thus  obtained  was  too  small  to  pay  for 
such  close  work.  Besides,  the  percentage  of  lead  in  the  ore  was 
reduced  only  0.5  per  cent. 

(c)  Iron  Pyrites. —  Either  intermixed  with  the  zinc  blende  or 
intersecting  the  same  in  narrow  streaks.     It  contains  about  $12 
gold  per  ton,  but  constitutes  only  a  comparatively  small  per- 
centage of  the  ore. 

(d)  Copper  Pyrites.  —  Occurs  seldom  and  then  only  in  very 
small  quantities. 

(e)  Native  Silver.  —  Now  and  then  specimens  with  metallic 
silver  in  flakes  or  wire  are  found. 

(/)  Gangue.  —  The  minerals  forming  the  gangue  are  quartz 
and  calcspar. 

The  ore  looks  like  solid  zinc  blende;  it  is  heavy  and  solid, 
showing  but  very  little  gangue.  In  the  following  table  two 
analyses  are  given.  Each  one  represents  the  average  of  large 
lots  of  ore.  The  one  is  unassorted,  just  as  it  was  extracted 
from  the  mine,  while  from  the  other  the  galena  and  gangue  were 
sorted. 

ANALYSIS  OF  SAN  FRANCISCO  DEL  ORO  ORE 


UNASSORTED 
ORE 

ASSORTED 
ORE 

Zinc 

2408 

2550 

Lead  . 

11  92 

11.56 

Iron  

700 

6.50 

Manganese.     .    . 

070 

0.53 

Cadmium  

0.16 

0.30 

Antimony  

0.50 

0.52 

CoDDer 

0  72 

1  02 

YYrI~*  

Alumina  . 

1  30 

365 

Calcium  carbonate 

982 

8.00 

Sulphur   ....        ... 

21  35 

21.01 

Nickel  

0.20 

Silver  

0.10 

0.12 

Gold  

trace 

trace 

Soluble  silica 

092 

Insoluble  gangue  

21.32 

19.41 

Total 

i  nn  no 

no  19 

Near  the  surface  the  ore  contained  more  free  galena  and  less 
zinc  blende;  and  it  is  said  that  the  first  owners  made  quite  a 
financial  success  by  smelting  the  ore  in  Mexican  furnaces.  How- 
ever, when  the  character  of  the  ore  changed,  the  mine  changed 


CHLORIDIZING  OF  ARGENTIFEROUS   ZINC^LEAD' 'ORE      101 

hands,  and  all  subsequent  attempts  to  work  the  ore  proved  a 
failure..  Though  the  ore  was  offered  much  cheaper  than  other 
ore  to  the  custom  mills  of  Parral,  only  small  quantities  were 
bought,  being  merely  used  as  flux  for  the  oxidized  ores  of  the 
Veta  Grande  and  other  mines  of  the  district  to  facilitate  chloridiz- 
ing  roasting.  Many  attempts,  however,  were  made  to  work  the 
ore  by  itself,  but  without  success.  Even  smelting,  which  ought 
to  have  been  out  of  the  question,  was  tried.  The  main  difficulty 
was  found  to  be  in  roasting.  The  ore  caked  very  readily  and  the 
silver  could  not  be  chloridized,  at  least  not  above  17  or  20  per 
cent.  After  many  unsuccessful  attempts,  further  trials  were 
abandoned  until  an  English  company  purchased  the  property. 

The  exceedingly  refractory  character  of  the  ore  and  the  very 
discouraging  experience  of  others  induced  the  managing  director 
of  the  English  company  to  have  elaborate  experiments  made 
before  erecting  a  mill  near  the  mine,  the  new  mill  to  be  constructed 
in  conformity  with  the  observations  and  experience  derived 
from  the  experiments.  For  this  purpose  the  Bosque  mill  at 
Parral,  an  old  25-stamp  lixiviating  mill,  was  purchased,  and  I  was 
commissioned  to  conduct  the  experiments. 

The  Mill.  —  It  consisted  of  25  stamps,  rock-breaker  and  self- 
feeders,  one  large-size  Stetefeldt  furnace,  claimed  to  roast  60 
tons  per  day,  one  brick-lined  revolving  cylinder  furnace  of  the 
White-Howell  type,  24  ft.  long  and  4  ft.  in  diameter,  and  two 
leaching  plants,  one  of  11  the  other  of  10  leaching  vats,  of  vari- 
ous sizes,  averaging  about  10  ft.  diameter  by  3  ft.  6  in.  depth, 
which,  for  want  of  grade,  were  almost  buried  in  the  ground. 
The  whole  arrangement  of  the  mill  was  ridiculously  inconvenient, 
causing  a  never-ending  handling  of  the  ore,  supplies  and  products. 
For  experimental  purposes,  however,  the  mill  was  good  enough, 
especially  as  I  had  the  privilege  of  erecting  any  other  roasting 
furnace  which  I  should  consider  advisable  to  experiment  with. 

ROASTING  EXPERIMENTS 

Theory.  —  As  shown  by  the  analysis,  the  ore  consisted  prin- 
cipally of  the  sulphides  of  zinc,  lead  and  iron.  The  other  sul- 
phureted  minerals  occurred  in  such  small  quantities  that  it  was 
not  necessary  to  pay  any  attention  to  the  chemical  part  they 
took  in  the  process  of  chloridizing  roasting. 


102  HYDROMETALLURGY  OF  SILVER 

Zinc  blende,  if  subjected  to  the  oxidizing  action  of  the  air, 
is  converted  into  zinc  oxide  and  zinc  sulphate,  while  sulphurous 
acid  escapes.  In  presence  of  salt,  zinc  sulphate  remains  indif- 
ferent, and  does  not  decompose  the  salt,  at  least  not  at  the  tem- 
perature used  in  chloridizing  roasting.  If  pure  zinc  blende,  finely 
pulverized,  and  mixed  with  salt,  is  placed  on  a  roasting  dish, 
and  roasted  in  the  muffle,  no  chlorine  gas  can  be  detected,  even 
if  exposed  to  a  bright  red  heat.  Even  freshly  prepared  zinc 
sulphate  mixed  with  salt  and  exposed  to  the  heat  of  the  muffle 
does  not  produce  any  chlorine  gas.  Zinc  blende,  therefore,  does 
not  take  an  active  part  in  producing  chlorine  during  chloridizing 
roasting,  at  least  not  enough  to  be  of  practical  value.  In  the 
roasted  ore  we  find,  therefore,  the  zinc  mostly  as  oxide  and  sul- 
phate and  so  also  in  the  flue-dust. 

If  galena  is  subjected  to  a  chloridizing  roasting,  especially  in 
presence  of  sufficient  air,  most  of  the  lead  is  converted  into  a  sul- 
phate, which,  as  such,  like  the  zinc  sulphate,  does  not  react  on 
the  salt,  and  therefore  does  not  generate  chlorine. 

Iron  pyrites  is  converted  into  ferric  oxide  and  into  ferrous 
and  ferric  sulphates,  both  of  which  react  very  energetically  on 
salt,  and  generate  chlorine. 

We  have,  therefore,  two  non-generators  and  only  one  generator 
of  chlorine  in  the  ore.  The  non-generators  of  chlorine,  galena 
and  zinc  blende,  however,  contain  all  the  silver,  while  the  iron 
pyrites  carries  only  some  gold,  but  no  silver.  This  is  a  very  im- 
portant point  to  take  into  consideration.  The  next  important 
point  is  the  fact  that  on  account  of  the  great  density  of  the  zinc 
blende  it  requires  a  long  time  to  oxidize.  Likewise  the  galena 
requires  long  roasting  at  a  low  heat,  while  iron  pyrites  decomposes 
quickly,  and  in  presence  of  salt  generates  chlorine  at  a  period  of 
the  roasting  when  neither  the  zinc  blende  nor  the  galena  are 
sufficiently  oxidized  to  yield  their  silver  to  the  action  of  the 
chlorine.  If,  therefore,  the  salt  is  mixed  with  the  ore  in  the 
stamp-battery,  the  chlorine  produced  by  the  reaction  of  ferric 
sulphate  and  salt  is  lost,  and  a  very  imperfect  chlorination  of  the 
silver  takes  place,  no  matter  how  long  roasting  may  be  continued 
and  how  much  salt  may  be  used.  For  instance,  when  the  ore 
was  roasted  with  12  per  cent,  of  salt  in  the  Stetefeldt  furnace, 
the  roasted  ore  contained  but  1.38  per  cent,  of  chlorine,  of  which 
0.8  per  cent,  was  combined  with  sodium,  which  represents  1.3  per 


CHLORIDIZING  OF  ARGENTIFEROUS   ZINC-LEAD  ORE       103 

cent,  undecomposed  sodium  chloride.  The  salt  used  in  roasting 
contained  but  78.25  per  cent,  of  sodium  chloride,  and  the  12 
per  cent,  of  salt  represents,  therefore,  9.39  per  cent.  Not  taking 
into  consideration  the  loss  in  weight  which  the  ore  sustained  in 
roasting,  we  can  assume  that  8.09  per  cent,  sodium  chloride  was 
decomposed  or  volatilized,  while  not  more  than  15  per  cent,  of 
the  silver  was  chloridized,  showing  that  the  silver  was  not  yet  in 
a  proper  condition  to  be  acted  upon  by  the  chlorine  at  the  time 
when  the  reaction  took  place  between  the  iron  sulphate  and  the 
salt.  Only  a  very  small  percentage  of  chlorine  was  found 
to  have  combined  with  other  bases.  The  chlorine  was  prac- 
tically an  entire  loss.  Similar  observations  were  made  by 
roasting  in  other  furnaces. 

In  roasting  an  ore  like  the  San  Francisco  del  Oro  ore,  it  is 
therefore  of  the  greatest  importance  to  add  the  salt  afterward 
and  not  to  mix  it  with  the  ore  in  the  stamp-battery.  But  this 
is  not  the  only  condition  to  be  observed.  We  have  to  take  into 
consideration  that  in  this  case  we  are  relying  on  the  sulphates  of 
iron  to  generate  chlorine,  and  that  these  sulphates  easily  decom- 
pose, forming  oxides.  If,  therefore,  the  oxidizing  period  is  con- 
tinued until  the  zinc  blende  and  galena  are  well  oxidized,  which 
takes  a  long  time,  we  will  have  no  iron  sulphate  left  to  decompose 
the  salt,  and  in  consequence  will  have  a  very  badly  chloridized 
ore.  At  a  high  temperature  these  iron  salts  decompose  more 
quickly,  giving  off  their  sulphuric  acid,  and  in  order  to  retain 
them  as  long  as  possible  the  ore  has  to  be  roasted  at  a  low  tem- 
perature. To  know  the  proper  time  when  the  salt  is  to  be  added 
is  of  the  greatest  importance;  this  knowledge,  however,  can  be 
obtained  only  by  repeated  tests  and  very  close  observation.  The 
most  suitable  time  to  add  the  salt  for  the  San  Francisco  del  Oro 
ore  was  found  to  be  when  the  black  color  of  the  ore  turns  brown 
but  still  shows  black  particles.  If  at  that  time  the  salt  is  added, 
a  distinct  odor  of  chlorine  can  be  observed,  which  lasts  during  the 
whole  of  the  finishing  period ;  while,  if  the  salt  is  added  too  soon 
or  too  late,  no  chlorine  evolves  from  the  ore  during  this  period. 

The  best  results  with  this  ore  could  undoubtedly  be  obtained 
by  subjecting  it  first  to  a  dead  oxidizing  roasting,  then  adding  a 
mixture  of  green  vitriol  and  salt.  But  the  ore  is  not  rich  enough 
to  permit  such  an  expense. 

While  the  incapacity  of  the  zinc  and  lead  sulphates  to  de- 


104  HYDROMETALLURGY  OF  SILVER 

compose  the  salt  makes  the  process  of  roasting  difficult  and 
complicated,  it  offers,  on  the  other  hand,  the  advantage  that  it 
reduces  materially  the  consumption  of  salt.  An  addition  of  four 
or  five  per  cent,  of  salt  gives  the  same  result  as  eight  or  ten  per 
cent.  If  the  roasting  is  very  carefully  conducted  even  three  per 
cent,  gives  good  results. 

Remarks.  —  The  most  difficult  and  at  the  same  time  the  most 
important  process  in  the  treatment  of  ores  by  wet  methods  is 
undoubtedly  the  chloridizing  roasting.  It  is  always  the  safest 
plan  for  the  operator  to  roast  as  thoroughly  as  possible.  If  the 
silver  is  well  chloridized,  the  sodium  hyposulphite  will  extract  all 
the  silver  chloride  and  frequently  will  leave  the  tailings  even 
poorer  than  indicated  by  the  chlorination  test,  without  the  use 
of  additional  solutions  or  chemicals,  thus  saving  time  and  expense, 
and  not  complicating  the  process.  A  high  chlorination  does  not 
necessarily  involve  a  high  loss  by  volatilization.  I  have 
recorded  instances  in  which  the  volatilization  of  silver  was  greater 
in  imperfectly  chloridized  charges  than  in  well-chloridized  ores 
(see  Chapter  IV,  "Loss  of  Silver  by  Volatilization").  Solutions,  by 
which  we  can  correct  a  badly  roasted  charge,  like  chloride  of 
copper  applied  during  base-metal  leaching,  or  by  which  part  of 
the  unchloridized  silver  can  be  extracted,  like  potassium  cyanide, 
or,  as  in  some  instances,  Russell's  extra  solution,  are  very  useful 
and  acceptable;  but  to  neglect  roasting  and  to  rely  for  closer 
extraction  on  these  solutions  is  a  rather  dangerous  practice. 

Being  convinced  that  the  successful  working  of  the  San 
Francisco  del  Oro  ore  depended  on  a  successful  roasting,  and 
knowing  the  great  difficulty  which  the  nature  of  this  ore  offered 
to  chloridizing  roasting,  particular  attention  was  paid  to  this  pro- 
cess, and  careful  studies  were  made  of  the  peculiarities  of  the  ore. 
The  principal  points  to  ascertain  were:  first,  the  mode  of  treat- 
ment which  the  ore  required  with  regard  to  temperature,  roasting 
time,  draft,  and  the  proper  time  for  adding  the  salt;  and,  second, 
which  of  the  furnaces  would  comply  best  with  the  requirements 
and  at  the  same  time  perform  the  work  the  cheapest. 

ROASTING  IN  THE  STETEFELDT  FURNACE 

The  furnace  was  a  large-size  Stetefeldt  furnace,  claimed  to  roast 
60  tons  per  day,  and  was  built  according  to  the  best  improved  design. 


CHLORIDIZING  OF  ARGENTIFEROUS   ZINC-LEAD  ORE       105 


Though  quite  elaborate  experiments  were  made,  it  was  not 
possible  to  obtain  good  results;  in  fact,  they  were  far  from  being 
satisfactory.  Nevertheless  they  are  interesting  and,  at  the  same 
time,  useful,  inasmuch  as  they  establish  the  fact  that  an  ore  like 
the  Del  Oro,  containing  25.5  per  cent,  zinc,  11.56  per  cent,  lead, 
and  21  per  cent,  sulphur,  is  by  far  too  refractory  for  the  Stetefeldt 
furnace.  Such  an  ore  requires  to  be  submitted  to  a  long  and 
gradually  increasing  temperature  before  the  salt  is  added.  The 
principle  of  the  Stetefeldt  furnace,  however,  is  just  the  reverse  of 
this  important  condition,  and  the  results,  as  a  matter  of  course, 
could  not  be  satisfactory. 

Notwithstanding  that  there  was  a  very  powerful  draft  through 
the  furnace,  and  that  the  ore  was  crushed  through  40-mesh 
screen,  only  about  three-eighths  of  the  ore  (by  volume)  was  carried 
by  the  draft  into  the  descending  flue  and  deposited  at  the  bottom 
of  the  chambers,  while  about  five-eighths  of  it  came  down  the 
shaft.  Actual  dust  was  not  carried  beyond  the  second  dust- 
chamber,  and  even  there  only  small  quantities  deposited,  owing 
to  the  great  specific  gravity  of  the  ore. 

After  heating  the  furnace  gradually  for  three  days,  charging 
was  begun.  There  were  20  stamps  running,  which  crushed 
through  40-mesh  screen  from  20  to  22  tons  per  twenty-four  hours, 
according  to  the  amount  of  salt  used.  The  experiments  were 
begun  with  an  addition  of  8  per  cent,  salt,  increasing  the  amount 
during  the  time  of  the  experiments  to  12  and  finally  to  16  per 
cent.  It  was  also  tried  to  roast  oxidizingly  and  to  add  the  salt  at 
intervals  to  the  ore  at  the  bottom  of  the  shaft,  but  without  success. 
The  furnace  was  kept  running  for  seven  days  and  was  then  stopped, 
as  no  signs  of  improvement  in  the  work  could  be  noticed  and  the 
cooling  floor  was  filled  with  badly  roasted  ore. 

Before  making  any  comments,  the  average  results  obtained 
under  different  conditions  are  given: 

ROASTING  WITH  8  PER  CENT.  SALT 


SHAFT 
Oz.  PER  TON 

DESCENDING  FLUE 
Oz.  PER  TON 

Average  of  .raw  ore  including  salt  .... 
Average  of  roasted  ore   . 

32.37 
34  41 

32.37 
2967 

Average  of  leach  tailings   

28  62 

23.72 

Average  of  chlorination 

16  90  per  cent 

20  20  per  cent. 

106 


HYDROMETALLURGY  OF  SILVER 


ROASTING  WITH  12  PER  CENT.  SALT 


SHAFT 
Oz.  PER  TON 

DESCENDING  FLUE 
Oz.  PER  TON 

Average  of  raw  ore  including  salt.  .  .  . 
Average  of  roasted  ore  
Average  of  leach  tailings  
Average  of  chlorination  

31.00 
31.33 
26.60 
15.20  per  cent. 

31.00 
27.50 
10.78 
60.80  per  cent. 

ROASTING  WITH  16  PER  CENT.  SALT 


SHAFT 
Oz.  PER  TON 

DESCENDING  FLUE 
Oz.  PER  TON 

Average  of  raw  ore  including  salt  .... 
Average  of  roasted  ore 

30.32 
33  39 

30.32 
2981 

Average  of  leach  tailings 

28.35 

22.30 

Average  of  chlorination  

15.20  per  cent. 

25.20  per  cent. 

An  experiment   was  also   made   merely  to   oxidize  the  ore 
without  adding  any  salt: 


OXIDIZING  ROASTING 


SHAFT 
Oz.  PER  TON 

DESCENDING  FLUE 
Oz.  PER  TON 

Average  of  roasted  ore 

3491 

28.57 

Average  of  leach  tailings 

32.65 

17.41 

Average  of  extractable  silver 

6.  50  per  cent. 

39.20  per  cent. 

The  oxidized  ore  from  the  descending  flue  was  charged  into  a 
tank,  treated  at  first  with  a  diluted  solution  of  cupric  chloride, 
then  leached  with  water  and  afterward  with  sodium  hyposul- 
phite, by  which  tailings  were  obtained  containing  14.43  oz.  per 
ton,  showing  an  extraction  of  49.5  per  cent,  silver. 

REROASTING  THE  ORE  FROM  THE  SHAFT 

The  ore  which  had  passed  through  the  Stetefeldt  furnace  once 
was  sifted,  to  free  it  from  lumps,  and  charged  a  second  time. 


CHLORIDIZING  OF  ARGENTIFEROUS   ZINC-LEAD  ORE       107 


SHAFT 
Oz.  PER  TON 

DESCENDING  FLUE 
Oz.  PER  TON 

Average  of  roasted  Ore 

31  49 

31  49 

Average  of  leach  tailings  
Average  of  chlorination  .  . 

29.60 
9.40  per  cent. 

24.79 
22.  30  per  cent. 

Observations  and  Comments.  —  If  we  compare  the  above 
results  we  find  that  those  obtained  in  the  shaft  were  pretty 
nearly  equally  bad,  whether  more  or  less  salt  was  used.  In  the 
descending  flue  12  per  cent,  of  salt  gave  the  best  result  (60.8  per 
cent,  chlorination).  An  excess  of  salt  lowered  the  chlorination. 
The  same  observation  was  made  afterward  in  the  Howell  and 
reverberatory  furnaces. 

The  ore  from  the  shaft  has  a  very  dark,  almost  black,  color, 
and  emits  volumes  of  sulphurous  acid  gas  when  discharged,  but  no 
chlorine.  By  leaving  the  ore  in  a  pile  on  the  cooling  floor  it 
continues  to  emit  sulphurous  acid  gas  for  a  couple  of  days,  without 
a  marked  increase  in  the  percentage  of  chloridized  silver  taking 
place.  In  dropping  through  the  shaft  the  main  portion  of  the 
ore  is  transformed  into  minute  globules,  which  show  that,  while 
the  ore  falls,  it  is  partially  slagged.  Trying  to  avoid  this,  the 
fire  was  lowered  so  much  that  the  lower  part  of  the  shaft  was 
quite  dark,  while  on  the  auxiliary  grate  the  fire  was  allowed  to  go 
out  entirely.  This,  however,  did  not  produce  any  change;  the 
ore  came  down  now  as  before  in  globules,  while  in  the  descending 
flue  the  temperature  continued  to  be  very  high.  It  is  apparent 
that  the  ore  when  sifted  into  the  shaft  creates  by  the  sudden 
combustion  of  the  sulphurets  an  extremely  high  temperature  in 
the  upper  regions  of  the  shaft,  which  causes  the  suspended  ore 
particles  to  melt  and  slag  to  globules.  Also  that  under  such 
circumstances  silicates  are  formed,  which,  incrusting  the  ore 
particles,  prevent  their  further  oxidation  and  chlorination. 

An  analysis  showed  that  the  roasted  ore  still  contained  8  per 
cent,  of  unoxidized  sulphur.  Ores  containing  not  more  than 
8  per  cent,  sulphur  usually  roast  well  in  a  Stetefeldt  furnace,  and 
it  was  expected  that  by  charging  the  ore  a  second  time  good 
results  might  be  obtained.  The  ore  was  sifted,  to  free  it  from 
lumps,  and  charged  again.  The  results,  however,  proved  to  be 
worse  than  those  obtained  in  the  first  roasting.  The  chlorination 


108  HYDROMETALLURGY  OF  SILVER 

of  15  and  16  per  cent,  was  reduced  to  9.4  per  cent.  The  ore 
maintained  its  dark  color  and  continued  to  emit  heavy  fumes  of 
sulphurous  acid.  It  is  not  probable  that  part  of  the  silver 
chloride  of  the  first  roasting  was  decomposed  by  passing  a  second 
time  through  the  furnace;  it  is  more  likely  that  the  dropping  of 
the  chlorination  in  the  shaft  was  caused  by  mechanical  separa- 
tion. Part  of  the  lighter  ore  particles  which  contained  less  lead 
and  were  better  chloridized  in  the  first  roasting  were  carried 
over  into  the  flue  during  the  second  roasting,  thus  seemingly 
reducing  the  original  percentage  of  chlorination.  Besides  part  of 
the  silver  chloride  may  have  slagged  during  the  second  roasting. 

There  are,  however,  other  circumstances  which  also  act  dis- 
advantageously.  For  instance,  it  was  found  after  the  first 
roasting  that  the  ore  in  the  shaft  contained  8.48  per  cent,  un- 
oxidized  sulphur  and  11.19  per  cent,  lead,  while  that  from  the 
descending  flue  contained  only  0.51  per  cent,  unoxidized  sulphur 
and  3.11  per  cent.  lead.  This  shows  that  a  separation  takes 
place,  the  shaft  receiving  the  main  portion  of  the  lead,  while  the 
lighter  minerals,  among  them  the  iron  pyrites,  are  carried  over 
into  the  descending  flue.  The  better  oxidation  and  chlorination, 
as  well  as  the  higher  temperature  in  the  flue,  are  principally  due 
to  this  separation.  But  as  the  main  bulk  of  the  ore  drops  through 
the  shaft  such  a  separation  is  disadvantageous. 

Another  interesting  fact  has  to  be  recorded.  Using  so  much 
salt,  and  obtaining  such  an  imperfect  oxidation  and  chlorination, 
we  should  naturally  expect  to  find  most  of  the  salt  undecomposed 
in  the  ore.  This,  however,  is  not  the  case.  In  the  shaft  the 
ore  contained  only  1.38  per  cent,  chlorine,  of  which  0.8  per  cent. 
was  combined  with  sodium,  representing  only  1.30  per  cent, 
undecomposed  sodium  chloride.  The  material  from  the  first 
dust-chamber  contained  0.38  per  cent,  of  chlorine,  of  which  0.16 
per  cent,  was  combined  with  sodium  representing  only  0.27  per 
cent,  sodium  chloride,  while  the  material  from  the  descending 
flue  contained  but  0.2  per  cent,  chlorine.  The  fine  white  dust 
from  the  last  dust-chamber  contained  0.42  per  cent,  chlorine  and 
a  great  deal  of  sulphuric  acid.  The  salt,  therefore,  was  decom- 
posed, and  the  chlorine,  either  as  hydrochloric  acid  or  as  chlorine, 
escaped  as  gas  without  effect.  If  it  were  volatilized  we  ought 
to  have  found  more  of  it  in  the  last  dust-chamber. 

Another  bad  feature  is  the  formation  of  lumps.     In  the  upper 


CHLORIDIZING  OF  ARGENTIFEROUS   ZINC-LEAD  ORE       109 

region  of  the  shaft,  the  ore-dust,  wherever  it  comes  in  contact 
with  the  hot  walls,  sticks  to  and  incrusts  them.  This  crust  peels 
off  and  drops  down  in  pieces.  It  is  almost  raw,  and  some  of  the 
larger  pieces,  when  broken,  show  the  texture  of  matte.  They 
form  in  large  quantities.  When  the  ore  was  sifted  for  reroasting, 
there  was  not  less  than  25  per  cent,  of  the  whole  ore  in  the  shape 
of  hard  lumps.  These  lumps  contained  only  21.87  oz.  silver  per 
ton,  and  0.82  per  cent,  chlorine,  of  which  0.28  per  cent,  was  combined 
with  sodium,  equal  to  0.46  per  cent,  sodium  chloride. 

Besides  the  chemical  difficulties,  a  very  annoying  mechanical 
difficulty  was  encountered.  The  ore-dust  in  the  same  way  as  it 
incrusted  the  walls  of  the  shaft  also  incrusted  the  lower  side  of 
the  screen  of  the  feeding  machine,  and  thus  stopped  up  the 
holes,  which  necessitated  a  frequent  changing  of  the  screen,  as 
often  as  twice  a  day. 

Heavily  sulphureted  ores,  especially  if  they  carry  zinc,  re- 
quire more  draft  in  roasting  than  less  sulphureted  ores.  This  is 
especially  the  case  with  the  Stetefeldt  furnace,  where  the  ore  is 
exposed  for  such  a  short  time  to  the  action  of  the  air  and 
heat.  In  compliance  with  this  theory  all  the  air-doors  with 
which  the  furnace  is  provided  were  opened,  but  the  result  did  not 
improve. 

REROASTING  THE  ORE  OF   THE  STETEFELDT    FURNACE    IN 
THE  MODIFIED  HOWELL  FURNACE 

The  partly  roasted  ore  from  the  Stetefeldt  furnace,  after 
sifting,  was  fed  into  the  Howell  furnace.  Having  previously 
ascertained  that  the  ore  contained  only  1.3  per  cent,  of  salt,  6 
per  cent,  and  sometimes  8  per  cent,  more  salt  were  added.  The 
rate  of  feeding  was  changed  from  time  to  time  in  order  to  test 
the  working  capacity  of  the  cylinder  for  this  ore;  thus  the  rate 
varied  from  5  to  9  tons  per  twenty-four  hours.  The  following 
results  are  the  averages  of  33  charges: 

Average  of  roasted  ore 31.42  oz.  silver  per  ton. 

Average  of  leach  tailings 17.55  oz.  silver  per  ton. 

Average  of  chlorination 44.20  per  cent. 

The  consumption  of  wood  in  reroasting  proved  to  be  much 
greater  than  in  roasting  the  raw  ore.  The  main  portion  of  the 
sulphur,  especially  that  of  the  pyrites,  which  easily  ignites,  hav- 


110  HYDROMETALLURGY  OF  SILVER 

ing  been  burnt  off  in  the  Stetefeldt  furnace,  the  ore  did  not  create 
any  heat  by  itself,  and  all  the  required  heat  had  to  be  furnished. 
To  reroast  8.3  tons  it  took  26  cargas  of  wood  (12  cargas  =  1  cord) 
=  0.27  cord  per  ton,  while  it  took  only  16  cargas  to  roast  10  or  11 
tons  of  raw  ore,  which  is  equivalent  to  0.13  cord  per  ton  of  ore. 
The  roasting  capacity  of  the  furnace  was  not  increased  by  roast- 
ing this  material,  which  contained  only  about  8  per  cent,  sulphur, 
but  on  the  contrary  was  diminished  as  compared  with  the  raw  ore 
containing  over  21  per  cent,  sulphur. 

The  reroasted  ore  was  of  a  red-brown  color,  smelled  of  chlorine, 
and  did  not  emit  any  sulphurous  acid  gas,  but  it  still  consisted 
principally  of  little  globules.  Quite  a  large  portion  of  the  globules 
remained  black,  no  matter  how  long  the  ore  was  kept  in  the  fur- 
nace. Some  of  them  were  magnetic,  but  the  great  majority 
were  not.  Between  the  fingers  the  reroasted  pulp  felt  sharp, 
like  pulverized  glass.  The  temperature  was  kept  at  a  proper 
degree,  and,  the  dust-chambers  and  furnace  having  been  pre- 
viously cleaned,  there  was  an  abundant  draft,  still  it  was  not 
possible  to  obtain  more  than  44.2  per  cent,  chlorination.  The 
cause  of  this  failure  wras  undoubtedly  the  silicates  which  were 
formed  during  the  roasting  in  the  Stetefeldt  furnace. 

APPLICATION  OF  STEAM 

A  jet  of  steam  was  introduced  into  the  reverberatory  of  the 
modified  Howell.  The  hydrochloric  acid  which  was  formed  by 
the  action  of  the  steam  was  of  decidedly  beneficial  influence,  and 
considerably  improved  the  result;  still  the  result  did  not  give 
entire  satisfaction.  The  same  percentage  of  salt  was  used  and 
the  same  temperature  maintained  as  in  the  foregoing  experi- 
ments, and  the  improved  results  are  therefore  exclusively  due  to 
the  action  of  the  hydrochloric  acid  on  the  silicates.  The  following 
figures  are  the  average  of  thirteen  charges: 

Average  of  reroasted  ore 31.00  oz.  per  ton. 

Average  of  leach  tailings 10.37  oz.  per  ton. 

Average  of  chlorination 66.60  per  cent. 

The  ore  still  contained  a  considerable  amount  of  these  little 
globules,  but  they  had  changed  their  color  to  red  brown,  and 
between  the  fingers  the  ore  felt  soft  and  not  so  sharp  and  glassy 
as  when  roasted  without  steam. 


CHLORIDIZING  OF  ARGENTIFEROUS   ZINC-LEAD  ORE       111 


CONCLUSIONS 

The  experiments  in  roasting  the  argentiferous  zinc  blende  and 
galena  ore  of  the  San  Francisco  del  Oro  mine  in  a  Stetefeldt  fur- 
nace have  shown: 

(1)  An  incomplete  oxidation  of  the  sulphureted  minerals,  the 
main  portion  of  the  ore  still  containing  8.48  per  cent,  unoxidized 
sulphur  when  roasted  with  salt,  and  7.6  per  cent,  when  roasted 
without  salt. 

(2)  An  insufficient  chlorination  of  the  silver.     The  highest 
chlorination  in  the  shaft  was  only  16.9  per  cent.,  and  as  62.5  per 
cent,  of  the  whole  volume  of  the  ore  dropped  into  the  shaft,  the 
somewhat   higher  chlorination  obtained   in  the  descending  flue 
could  not  much  improve  the  average  chlorination. 

(3)  That  the  principle  of  the  Stetefeldt  furnace  is  contrary 
to  the  conditions,  of  which  the  maintenance  is  so  essential  to 
roasting  ores  containing  much  zinc  blende  and  galena.     Instead 
of  permitting  the  ore  to  be  subjected  for  a  longer  time  at  a  low 
but  gradually  increasing  temperature,  the  ore,  entering  the  fur- 
nace, encounters  immediately  the  highest  temperature,  which  is 
detrimental  to  the  roasting  of  such  ores. 

(4)  That  on  account  of  the  sudden  exposure  of  the  raw  ore 
particles  to  such  a  high  temperature  they  melt  to  minute  globules, 
which  makes  the  ore  unfit  for  further  treatment. 

(5)  That  a  concentration  of  the  lead  minerals  takes  place  in 
the  shaft,  which  is  disadvantageous. 

(6)  That  about  25  per  cent,  of  the  ore  when  passing  through 
the  furnace  is  changed  into  hard  lumps  of  almost  raw  ore,  and 
that  the  construction  of  the  furnace  does  not  offer  any  means  to 
prevent  it. 

(7)  That  the  lower  side  of  the  feeding  screen  becomes  rapidly 
incrusted  and  the  holes  obstructed,  requiring  a  too  frequent  ex- 
change of  screens. 

These  observations  taken  together  prove  beyond  doubt  that 
the  Stetefeldt  furnace  is  not  suitable  for  the  San  Francisco  del 
Oro  ore,  and  consequently  for  no  ores  heavily  charged  with  zinc 
blende  and  galena. 


112  HYDROMETALLURGY  OF  SILVER 


ROASTING  IN  THE  WHITE-HOWELL  FURNACE 

The  furnace  which  was  used  was  not  a  regular  Howell.  It 
consisted  of  a  revolving  cylinder  of  uniform  diameter,  the  shell 
being  made  of  boiler  iron  and  lined  with  bricks  the  whole  length. 
The  principle  on  which  it  works,  however,  is  identical  with  that 
of  the  Howell,  and  this  name  is  used  here  merely  to  indicate  the 
type  of  furnace. 

The  Howell,  like  the  Stetefeldt  furnace,  requires  the  salt  to 
be  added  to  the  ore  before  entering  the  furnace.  With  some  ores 
this  is  immaterial,  but  it  is  a  point  of  the  greatest  importance 
for  the  Del  Oro  ore.  If  the  salt  is  previously  added  the  ore 
becomes  sticky,  incrusts  the  furnace  rapidly,  and  when  it  leaves 
the  furnace  consists  mostly  of  lumps,  and  what  is  still  worse, 
without  being  chloridized.  If  the  ore  is  charged  without  salt  it 
remains  dry  and  sandy,  but  a  very  annoying  separation  takes 
place.  The  fine  particles  are  carried  by  the  draft  into  the  dust- 
chambers,  and  only  the  coarse  sand  passes  through  the  furnace, 
without  being  sufficiently  desulphurized.  If,  then,  salt  is  added 
in  the  drop-pit,  only  a  small  percentage  of  the  silver  becomes 
chloridized;  the  best  results  gave  only  29  per  cent,  chlorination. 

In  order  to  diminish  the  separation,  two  per  cent,  of  salt  was 
added  to  the  ore  in  the  battery,  while  the  balance  of  the  salt 
was  added  in  the  drop-pit.  This  small  percentage  of  salt  made 
the  ore  sticky  enough  to  diminish  considerably  the  dusting, 
without  causing  the  formation  of  lumps  or  too  heavy  an  incrus- 
tation of  the  furnace. 

By  this  mode  of  roasting  the  chlorination  improved  consider- 
ably, the  average  of  three  days'  run  being  67  per  cent.  It  was 
soon  evident,  however,  that  the  Howell  furnace,  as  such,  could  not 
roast  the  Del  Oro  ore.  The  results  were  not  uniform  and  relia- 
ble, being  sometimes  high,  sometimes  low;  and  notwithstanding 
the  greatest  care  the  average  could  not  be  brought  above  67  per 
cent.  But  the  roasted  ore  was  in  a  good  condition;  it  was  un- 
finished but  not  spoiled,  as  in  the  Stetefeldt  furnace,  and  there 
was  reason  to  expect  that,  by  an  alteration  which  would  give 
the  ore  more  roasting  time  and  allow  a  better  regulation  of  the 
temperature,  good  results  would  be  obtainable. 


CHLORIDIZING  OF  ARGENTIFEROUS   ZINC-LEAD   ORE       113 


ROASTING  IN  THE  MODIFIED  HOWELL  FURNACE 

In  front  of  the  furnace  I  constructed  a  shallow  drop-pit  and 
a  fireplace,  the  long  side  of  the  fireplace  being  opposite  the  dis- 
charge of  the  furnace,  so  that  the  flame  before  entering  the 
furnace  had  to  traverse  the  drop-pit.  To  one  side  of  the  drop-pit, 
and  communicating  with  it,  there  was  attached  a  small  reverber- 
atory  furnace  6  x  8  ft.,  the  bottom  of  both  being  on  the  same 
level.  The  reverberatory  contained  one  working  door  and  a 
24-in.  fireplace.  When  enough  ore  had  accumulated  in  the  pit 
to  make  a  charge  for  the  reverberatory,  it  was  pushed  by  means 
of  a  hoe  into  the  reverberatory.  Each  charge  consisted  of  about 
1400  Ib.  While  starting  the  furnace  a  strong  fire  was  kept  in 
both  fireplaces,  but  after  the  process  was  in  operation  the  fire  in 
front  of  the  cylinder  was  much  lowered;  in  fact,  so  much  so  that 
half  the  grate-bars  remained  bare  of  wood.  Only  now  and  then 
a  thin  stick  of  wood  was  added,  just  enough  to  prevent  the  drop- 
pit  from  getting  chilled.  Two  per  cent,  of  salt  was  added  to  the 
ore  in  the  batte^. 

If  the  roasting  is  properly  conducted,  the  blue  flame  of  the 
ignited  pyrites  can  be  observed  in  the  back  part  of  the  cylinder. 
Next  to  it  and  reaching  beyond  the  middle  of  the  cylinder  the  ore 
assumes  a  higher  temperature,  forming  a  belt  of  bright-red  heat. 
In  this  region  the  principal  oxidation  takes  place,  and  the  increase 
in  temperature  is  caused  by  the  oxidation  and  not  by  its  position 
nearer  to  the  fire.  The  part  of  the  furnace  next  to  the  fire  and 
nearly  one-third  of  the  whole  length  ought  to  look  dark.  The 
furnace  is  mostly  heated  by  the  combustion  of  the  sulphides, 
and  receives  but  little  supply  from  the  fireplace  and  from  the 
reverberatory.  In  fact,  the  ore  while  in  the  cylinder  should  be 
left  as  much  as  possible  to  roast  in  its  own  heat.  This  is  a  very 
important  condition  to  maintain.  The  object  is  to  convert  as 
much  as  possible  of  the  galena  and  zinc  blende  into  sulphates 
and  oxides  before  generating  chlorine,  and  to  avoid  until  then  as 
much  as  possible  the  decomposition  of  the  iron  salts.  This  can 
only  be  done  by  maintaining  a  low  heat  after  the  combustion  of 
the  pyrites.  An  excess  of  heat  is  invariably  connected  with  an 
excessive  loss  of  silver  by  volatilization  and  by  a  low  chlorination. 
Galena  and  zinc  blende  roast  quicker  and  better  in  a  low  than  in 


114  HYDROMETALLURGY  OF  SILVER 

a  high  heat.  When  the  ore  leaves  the  cylinder  and  drops  into 
the  pit  it  should  be  of  a  very  dull  red  heat,  while  the  color  after 
cooling  should  be  dark  yellow-brown. 

If  the  temperature  is  so  kept,  neither  the  odor  of  chlorine  nor 
much  of  sulphurous  acid  can  be  detected.  At  an  increased  heat, 
sulphurous  acid  emits  again  strongly,  showing  that  the  oxidation 
is  not  yet  completed.  As  the  temperature  in  the  cylinder  is 
mostly  produced  by  the  combustion  of  the  sulphureted  material, 
the  main  means  of  regulating  the  same  is  the  feed.  If  too  much 
ore  enters  the  furnace  the  belt  of  bright-red  heat  increases,  ad- 
vancing more  and  more  toward  the  front,  and  finally  the  whole 
furnace  assumes  this  temperature.  The  ore  dropping  into  the 
pit  is  very  hot,  emits  heavy  fumes,  and  overheats  the  pit.  If 
then  removed  into  the  reverberatory,  it  takes  a  very  long  time 
to  be  finished,  necessitating  an  interruption  in  the  feed  of  the 
cylinder.  On  the  other  hand,  if  insufficient  ore  is  charged,  the 
belt  of  bright-red  heat  gets  smaller  and  moves  toward  the  back 
end  of  the  furnace. 

When  the  properly  prepared  ore  enters  the  reverberatory 
furnace,  the  salt  is  added  and  the  temperature  is  somewhat 
increased.  It  commences  to  fume,  and  swells,  without  forming 
more  lumps  than  an  ordinary  ore.  In  the  beginning  strong 
fumes  of  sulphurous  acid  emit,  but  soon  cease,  and  chlorine 
appears.  The  charge  is  finished  if  the  fumes  assume  a  mild  and 
sweetish  smell  of  chlorine;  as  long  as  they  smell  strong,  roasting 
has  to  be  continued. 

A  series  of  experiments  was  made  to  ascertain  the  smallest 
amount  of  salt  practicable,  and  it  was  found  that  4,  6,  8,  and  10  per 
cent,  give  about  equal  results.  Twelve  per  cent,  begins  to  make 
the  ore  too  sticky  and  produces  less  chlorination;  3  per  cent,  is 
sufficient  if  the  roasting  is  very  carefully  conducted,  but  then 
only  1  per  cent,  has  to  be  added  in  the  battery  and  2  per  cent,  in 
the  furnace.  Four  per  cent.,  however,  is  safer,  as  then  the  result 
does  not  depend  so  much  on  the  skill  and  good-will  of  the  laborers. 

The  roasting  capacity  of  the  furnace  proved  to  be  much  less 
for  this  ore  than  for  ordinary  ore.  Not  more  than  8J  tons  could 
be  roasted  in  twenty-four  hours.  It  is  true  the  cylinder  was  only 
24  ft.  long,  but  even  with  a  32-ft.  cylinder  it  cannot  be  expected 
to  roast  more  than  12  tons  per  day.  Each  charge  had  to  remain 
two  hours  on  the  reverberatory  hearth.  Though  the  ore  was 


CHLORIDIZING  OF  ARGENTIFEROUS  ZINC-LEAD  ORE      115 

roasted  in  the  reverberatory  at  a  somewhat  increased  heat,  yet 
the  temperature  could  not  be  increased  beyond  dull  red  without 
losing  too  much  silver  by  volatilization. 

ADDITIONAL    CHLORINATION    AFTER    THE   ORE   HAS   LEFT    THE 

FURNACE 

Some  ores  gain  much  in  chlorination  if  left  hot  in  a  pile  for 
some  time.  This  is  mostly  the  case  when  an  ore  is  insufficiently 
roasted,  or  when  the  nature  of  the  ore  is  such  as  to  require  a  long 
roasting  time  at  a  low  heat.  Another  additional  chlorination 
can  be  produced  by  moistening  the  ore  and  leaving  it  for  several 
hours  in  a  pile.  This  is  usually  the  case  if  the  ore  contains  copper. 
Roasted  ore  containing  caustic  lime  should  not  be  left  moist  on 
the  cooling  floor.  The  most  important  additional  chlorination, 
however,  takes  place,  according  to  numerous  observations  of 
mine,  during  base-metal  leaching.  The  roasted  Del  Oro  ore 
either  contained  much  caustic  lime  —  which,  however,  is  hardly 
possible,  as  the  raw  ore  is  so  rich  in  sulphur  —  or  some  other  chemi- 
cal substance  which  acted  decomposingly  on  the  silver  chloride, 
because  it  could  not  be  moistened  on  the  cooling  floor  without 
sustaining  quite  a  loss  in  chlorination.  To  prevent  this,  the  vats 
were  charged  about  one-third  with  water  and  the  hot  ore  dumped 
into  it,  thus  producing  a  hot  base-metal  solution.  The  observa- 
tion was  made  that  by  this  practice  not  only  was  the.  decompo- 
sition of  the  silver  chloride  avoided,  but  that  a  considerable 
increase  in  the  silver  chlorination  took  place,  in  some  instances 
as  much  as  12.9  per  cent,  (see  following  table,  charge  No.  11). 

If  the  original  chlorination,  however,  was  75  per  cent,  or  more, 
this  additional  chlorination  amounted  to  much  less.  By  adding 
some  cupric  chloride  to  the  water  in  the  vat  before  dumping  the 
ore,  I  found  that  badly  roasted  charges  gained  in  chlorination  as 
much  as  34  to  38  per  cent,  (see  table,  charges  Nos.  9,  15  and  16). 

These  are  very  important  observations,  and  give  the  operator 
the  means  of  correcting  badly  roasted  charges. 

RESULTS 

The  table  on  page  1 16  is  a  record  of  the  results  obtained  in  roast- 
ing the  San  Francisco  del  Oro  ore  in  the  modified  Howell  furnace. 
It  represents  a  two  weeks'  run.  As  each  tank  charge  contained 


116 


HYDROMETALLURGY   OF  SILVER 


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CHLORIDIZING  OF  ARGENTIFEROUS  ZINC-LEAD  ORE     117 

the  whole  ore  of  twenty-four  hours'  roasting,  it  offered  a  good 
opportunity  to  follow  each  charge  from  the  raw  ore  down  to  the 
tailings,  and  to  ascertain  for  each  charge  the  loss  by  volatiliza- 
tion, gain  by  additional  chlorination,  etc. 

Taking  the  averages  of  the  results,  we  find  the  silver  chlori- 
nation when  the  ore  left  the  furnace,  68.4  per  cent.;  additional 
chlorination,  13.3  per  cent,  or  a  total  chlorination  of  81.7  per 
cent.  The  low  average  chlorination  of  the  ore  when  leaving  the 
furnace  was  caused  by  the  three  badly  roasted  charges  —  9,  15 
and  16.  The  other  eleven  charges  gave  an  average  of  about  75 
per  cent. 

The  total  or  actual  chlorination  of  81.7  per  cent,  may  seem 
to  be  low,  but  if  we  consider  that  the  ore  is  of  low  grade,  averag- 
ing only  28.8  oz.  per  ton,  and  that  1  per  cent,  represents  only 
0.28  of  an  ounce  silver;  and  also  consider  that  the  ore  contains 
about  37  per  cent,  zinc  blende  and  13  to  19£  per  cent,  galena, 
which  carry  all  the  silver,  and  that  the  ore  was  pronounced  as 
being  too  refractory  for  chloridizing  roasting,  we  have  to  count 
the  work  done  by  the  modified  Howell  furnace  as  very  satisfac- 
tory, especially  as  such  a  chlorination  secured  the  success  of  the 
enterprise  owing  to  the  cheapness  of  mining  and  reduction. 

Loss  OF  SILVER  BY  VOLATILIZATION 

The  loss  of  silver  by  volatilization  in  these  experiments  was 
ascertained  by  the  method  described  in  a  previous  chapter. 

Owing  to  the  fact  that  a  great  portion  of  the  lead  and  zinc 
sulphides  is  converted  into  sulphates,  the  San  Francisco  del  Oro 
ore  loses  but  a  small  percentage  of  its  weight  during  roasting. 
The  tests  showed  a  loss  of  2^  and  3J  per  cent. 

With  these  figures,  and  the  assay  value  of  the  raw  and  roasted 
ore,  the  loss  of  silver  by  volatilization  was  calculated.  The 
extremes  were  1.3  per  cent,  and  15.5  per  cent,  while  the  average 
gave  7.9  per  cent.  The  figures  contained  in  the  corresponding 
column  of  the  table  illustrate  how  variable  this  loss  is,  and  what 
a  severe  loss  of  silver  can  be  caused  by  even  slight  oversights. 
I  found  the  Del  Oro  ore  to  be  more  sensitively  disposed  for  such 
loss  than  many  others,  even  antimonial  ores  which  I  had  treated 
before.  The  least  increase  of  the  temperature  above  dull  red 
causes  a  heavy  loss,  even  if  this  increase  of  the  temperature  lasts 


118  HYDROMETALLURGY  OF  SILVER 

only  a  very  short  time.  Thus  two  or  three  thin  sticks  of  wood, 
if  thrown  on  the  fire  before  needed,  may  materially  increase  the 
loss. 

The  loss  by  volatilization  is  not  in  direct  proportion  to  the 
per  cent,  of  chloridized  silver.  Frequently  a  well  chloridized 
ore  suffers  much  less  loss  than  a  badly  chloridized  one.  Refer- 
ring to  the  table  we  find,  for  instance,  in  charge  No.  16  the  silver 
was  chloridized  only  to  47.2  per  cent,  while  the  loss  by  volatiliza- 
tion was  as  high  as  13  per  cent.  Again,  in  charge  No.  20,  76  per 
cent,  of  the  silver  was  chloridized,  while  the  loss  by  volatilization 
amounted  to  only  1.3  per  cent.  We  find  the  same  in  charges 
Nos.  12,  14,  etc. 

THE  ROASTED  ORE 

The  roasted  ore  contains  only  a  small  percentage  of  lumps. 
These  are  not  hard,  but  porous,  and  fall  to  powder  if  kept  in  con- 
tact with  water  for  some  time.  If  the  ore  is  left  dry  in  a  pile  it 
hardens.  If  left  undisturbed  for  a  week  or  two  it  becomes  so 
hard  that  it  requires  the  use  of  a  pick  to  loosen  it.  In  water, 
however,  it  softens  easily  again.  The  color  is  usually  red-brown, 
but  occasionally,  if  there  is  less  iron  pyrites  in  the  ore,  it  is  yellow- 
brown. 

The  analysis  of  the  unassorted  ore  after  roasting  (see  above 
analysis  of  the  raw  ore)  is  here  given.  The  ore  was  roasted  with 
5  per  cent,  of  salt. 

ANALYSIS  OF  UNASSORTED  ROASTED  ORE 

Gold trace. 

Silver 0.09 

Lead 9.00 

Iron 6.00 

Zinc 22.45 

Lime  (calculated  as  caustic) 5.65 

Antimony 0.75 

Copper 0.60 

Cadmium 0.10 

Alumina 3.09 

Soda  (calculated  as  caustic) 3.79 

Sulphuric  acid 13.16 

Chlorine 0.88 

Soluble  silica 8.00 

Insoluble  gangue 18.61 

Oxygen  of  the  oxides ? 

This  analysis  shows  that  the  heavy  metals  were  principally 
converted  into  sulphates,  and  that  only  a  small  portion  of  them, 


CHLORIDIZING  OF  ARGENTIFEROUS  ZINC-LEAD  ORE     119 

if    any,  can    be   present    as    chlorides.     The   0.88    per   cent,  of 
chlorine  may  be  due  to  undecomposed  salt. 

If  the  furnace  crust  is  not  removed  from  the  furnace  for  some 
time,  it  changes  its  color.  In  some  parts  it  is  greenish  white;  in 
others,  flesh-colored.  It  gets  very  hard,  and  when  moistened 
with  water  generates  heat  and  slacks  like  lime.  I  had  never 
made  this  observation  before,  not  even  with  the  very  calcareous 
ore  of  Las  Yedras.  It  is  to  be  regretted  that  it  was  overlooked 
to  make  an  analysis  of  this  crust.  It  is  difficult  to  believe  that 
this  phenomenon  could  be  caused  by  caustic  lime,  because  the 
latter  could  not  well  exist  in  an  atmosphere  of  sulphuric  and 
sulphurous  acid  and  of  chlorine. 

CONSUMPTION  OF  WOOD 

Owing  to  the  large  quantity  of  sulphureted  minerals  in  the 
ore,  and  the  very  low  temperature  at  which  the  Del  Oro  ore  has 
to  be  roasted,  the  consumption  of  wood  is  very  small.  After 
the  furnace  is  heated  and  the  cylinder  incrusted,  it  takes  hardly 
any  fire  in  front  of  the  cylinder  to  maintain  the  proper  tempera- 
ture. In  the  reverberatory  addition  a  little  more  fire  is  needed, 
but  much  less  than  ordinary  ores  require. 

During  two  weeks  the  wood  was  weighed  (it  was  bought  by 
weight)  and  the  total  consumption  during  this  time  was  found  to 
be  220  cargas  of  300  Ib.  With  this  amount  of  wood  115.8  tons 
of  ore  were  roasted,  which  gives  1.8  cargas  per  ton  of  ore. 
Twelve  cargas  of  the  Parral  wood  are  equal  to  one  cord,  and 
if  we  express  the  consumption  in  cords  we  find  that  with  one 
cord  of  wood  the  furnace  roasted  6.3  tons  of  ore. 

COST  OF  ROASTING  IN  THE  MODIFIED  HOWELL  FURNACE 

The  cost  of  roasting  8J  tons  per  twenty-four  hours  was  as 
follows  : 

Labor.  .  ....................................  $6.60 

4  per  cent,  salt,  680  Ib.  at  1.27?  ...............  8.63 

15.7  cargas  wood  at  75^  .......................  11.77 

Steam  power,  10  cargas  wood  at  75^  ............  7.50 

Oil,  light,  tools,  etc  ...........................  2.00 

Management,  office,  mechanic's  assay  office  ......  1.78 

8.5  =  S4.50 


Cost  per  ton  ............................  $4.50  Mexican  currency. 


120  HYDROMETALLURGY   OF  SILVER 

To  ascertain  the  cost  of  steam  power  a  separate  boiler  was 
used  for  the  furnace.  It  is  apparent  that  by  using  a  boiler  for 
only  one  small  furnace  the  expense  per  ton  of  ore  will  be  much 
greater  than  if  with  the  same  boiler  and  engine  several  large 
furnaces  are  operated.  But  this  was  the  only  way  of  getting  an 
estimate.  The  steam  for  working  the  pumps  and  preparing  the 
calcium  sulphide  was  supplied  by  the  same  boiler,  and  had  to  be 
charged  to  roasting. 

As  the  statement  of  cost  is  made  only  8J  tons  per  day,  it 
would  be  misleading  if  the  whole  expenses  for  management, 
mechanic's  assay  office,  etc.,  should  be  charged  to  the  8J  tons, 
as  those  expenses  will  be  about  the  same  for  100  tons  per  day, 
the  intended  capacity  of  the  new  mill.  The  expenses  for  manage- 
ment, etc.,  were,  therefore,  calculated  for  100  tons  per  day  and 
the  8J  tons  charged  in  proportion.  But  there  are  three  depart- 
ments in  the  mill,  viz.,  stamping,  roasting,  and  leaching,  and 
each  department  has  to  be  charged  with  one-third  of  this  expense. 
The  above  item  of  $1.78  represents,  therefore,  one-third.  In  the 
statement  of  cost  4  per  cent,  salt  was  put  down  because  subse- 
quent experiments  proved  this  amount  to  be  sufficient. 

SUMMARY 

It  must  be  borne  in  mind  that  the  figures  contained  in  the 
above  table  are  the  results  of  experiments.  In  other  words,  these 
figures  were  obtained  under  different  treatments  with  regard  to 
salt,  temperature,  time,  etc.,  and  the  averages,  therefore,  do  not 
represent  the  best  obtainable  results.  This  is  especially  the  case 
with  the  loss  by  volatilization,  which  will  be  less  after  the  men 
acquire  more  skill  in  maintaining  the  proper  temperature.  But 
as  the  roasting  results  obtained  with  the  modified  Howell  furnace 
under  above  conditions  are  good  enough  to  secure  a  profitable 
reduction  of  the  ore,  we  may  just  as  well  accept  these  averages 
as  a  basis  for  calculations  and  estimates. 

RECAPITULATIONS 

Average  value  of  raw  ore  without  salt 28.85  oz.  per  ton. 

Average  value  of  raw  ore  containing  salt 27.49  oz.  per  ton. 

Average  value  of  roasted  ore 26.10  oz.  per  ton. 

Average  value!  of  vat  tailings 4.76  oz.  per  ton. 

Average  per  cent,  of  chlorination 81.6  per  cent. 

Average  per  cent,  of  actual  extraction 74.9  per  cent. 


CHLORIDIZING  OF  ARGENTIFEROUS  ZINC-LEAD  ORE     121 

Average  number  of  ounces  silver  extracted  per 

ton 21.9  ounces. 

Average  percental  loss  by  tailings 17.0  per  cent. 

Average  percental  loss  by  volatilization 7.9  per  cent. 

Average  per  cent,  of  salt  used 4.7  per  cent. 

Average  number  of  tons  roasted  per  day  with 

one  furnace 8.5  tons. 

Average  cost  of  roasting  one  ton  of  ore $4.50  Mexican  currency. 

Average  consumption  of  wood  per  ton,  includ- 
ing steam  power  3  cargas. 

ROASTING  IN  THE  REVERBERATORY  FURNACE 

The  appliances  for  roasting  consisted  of  a  large  size  Stetefeldt 
furnace  and  one  24-ft.  revolving  cylinder  furnace,  as  described 
above.  After  the  Stetefeldt  furnace  proved  to  be  a  failure  with 
the  Del  Oro  ore,  the  roasting  capacity  became  reduced  to  that  of 
the  revolving  cylinder,  or  8J-  tons.  In  order  to  increase  the 
roasting  capacity  to  the  stamping  capacity,  and  for  the  sake  of 
further  experiments,  four  two-story  reverberatory  furnaces  were 
erected,  the  lower  hearth  of  220  sq.  ft.  surface  and  the  upper  of 
210  sq.  ft.  surface  (Figs.  10,  11,  and  12). 

Each  two-story  furnace  took  four  charges  of  one  ton  each. 
When  one  charge  was  finished  all  the  others  were  moved  forward, 
and  on  the  first  hearth  a  new  charge  dropped  through  an  opening 
in  the  arch.  From  the  second  hearth  the  charge  was  dropped  on 
the  lower  hearth,  through  an  opening  in  the  bottom  near  the 
working  door.  The  upper  hearth  was  used  exclusively  for  oxi- 
dizing roasting,  and  the  4  per  cent,  of  salt  for  better  mixing  was 
added  while  the  charge  was  dropped  on  the  lower  hearth.  As 
the  proper  time  when  the  salt  is  to  be  added  had  proved  to  be  a 
factor  of  great  importance  in  roasting  this  ore,  the  process  was 
conducted  on  the  appearance  of  the  ore  at  the  second  hearth. 
When  the  charge  on  that  hearth  showed  it  to  be  in  proper  condition 
to  receive  the  salt,  the  ore  from  the  finishing  hearth  was  dis- 
charged, the  charge  from  the  third  hearth  moved  on  the  finishing 
hearth  and  the  charge  on  the  second  hearth  dropped  on  the 
third,  while  the  salt  was  added.  Every  two  and  one-half  to  three 
hours  a  charge  was  done,  and  therefore  each  two-story  furnace 
roasted  from  8  to  nearly  10  tons  per  twenty-four  hours,  according 
to  the  quantity  of  lead  contained  in  the  ore.  Zinc  blende  roasts 
quicker  than  galena.  Each  charge  was  ten  to  twelve  hours  in 
the  furnace. 

The  chlorination  results  were  so  near  those  obtained  in  the 


122  HYDROMETALLURGY   OF  SILVER 

modified  Howell  furnace  that  no  details  need  to  be  given  here, 
but  details  and  observations  will  be  given  of  a  charge  which  was 
subjected  to  a  prolonged  oxidizing  roasting,  because  they  are 
rather  interesting.  A  charge  of  the  Del  Oro  ore  was  placed  directly 
on  the  hearth  nearest  to  the  fireplace  (finishing  hearth)  and  kept 
there  until  finished. 


OXIDIZING  ROASTING 

First  hour.  —  Assay  of  raw  ore,  35.57  oz.  per  ton. 

During  this  hour  the  ore  had  just  fairly  started  to  ignite.  A 
sample  when  cold  had  not  changed  its  color,  and  looked  like  raw 
ore.  In  this  and  the  following  hours  a  sample  was  leached  first 
with  water,  then  with  sodium  hyposulphite,  and  to  each  filtrate 
calcium  sulphide  was  added. 

Wash- water:  no  precipitate. 

Hypo-solution  (1  per  cent.) :  no  precipitate,  light  coloration. 

Concentrated  hypo:  a  heavy  precipitate,  consisting  mostly 
of  iron  and  lead. 

Assay  of  ore,  33.54  oz.  per  ton;  after  leaching  with  hypo,  33.54 
oz.  per  ton;  no  soluble  silver. 

Second  hour.  —  Toward  the  end  of  this  hour  the  ore  com- 
menced to  lose  its  own  heat,  caused  by  the  combustion  of  the 
pyrites.  During  this  hour  no  fire  was  kept  up.  The  color  of  the 
sample  when  cold  was  a  dark  greenish-brown. 

Wash- water:  no  precipitate  or  discoloration,  therefore  no 
salts  soluble  in  water. 

Hypo-solution  (1  per  cent.):  a  slight  precipitate. 

Hypo-solution  (3  per  cent.):  a  heavy  precipitate. 

Assay  of  ore,  34.41  oz.  per  ton;  after  leaching,  34.12  oz.  per 
ton;  no  soluble  silver. 

Third  hour.  —  During  this  hour  the  ore  had  lost  its  own  heat, 
and  fire  was  started  again,  but  the  temperature  was  kept  very 
low.  The  color  of  the  sample  when  cold  was  brown-yellow, 
more  brown  than  yellow. 

Wash- water:  no  precipitate,  no  coloration,  no  soluble  salts. 

Hypo-solution  (1  per  cent.):  considerable  precipitate. 

Assay  of  ore,  36.74  oz.  per  ton;  after  leaching,  35.28  oz.  per 
ton;  soluble  silver,  1.46  oz.  per  ton,  or  3.9  per  cent. 

Fourth   hour.  —  The   temperature   was    somewhat    increased 


CHLORIDIZING  OF  ARGENTIFEROUS  ZINC-LEAD  ORE     123 

during  this  hour,  but  still  kept  rather  low.  The  color  of  the 
sample  when  cold  was  of  a  much  lighter  yellowish  brown. 

Wash-water:  the  first  indication  of  precipitate,  showing  that 
until  the  end  of  the  fourth  hour  no  zinc  sulphate  had  been  formed. 

Hypo-solution  (1  per  cent.):  considerable  precipitate. 

Assay  of  ore,  35.57  oz.  per  ton;  after  leaching,  31.78  oz.  per 
ton;  soluble  silver,  3.79  oz.  or  10.6  per  cent. 

Fifth  hour.  —  The  temperature  still  kept  at  dull  red.  Color 
of  sample  when  cold  brown-red,  showing  that  at  this  time  some 
oxide  of  iron  had  formed. 

Wash- water:  considerable  precipitate  of  a  yellowish-white 
color,  mostly  zinc  and  cadmium. 

Assay  of  ore,  35.13  oz.  per  ton;  after  leaching,  22.89  oz.  per 
ton;  soluble  silver,  12.24  oz.  per  ton,  or  34.8  per  cent. 

Sixth  hour.  —  The  same  low  heat;  color  of  sample  when  cold 
darker  red-brown. 

Assay  of  ore,  33.83  oz.  per  ton;  after  leaching,  20.41  oz.  per 
ton;  soluble  silver,  13.42  oz.  per  ton,  or  39.6  per  cent. 

Seventh  hour.  —  The  same  temperature;  sample  when  cold 
still  darker  red-brown. 

Assay  of  ore,  30.74  oz.  per  ton;  after  leaching,  13.41  oz.  per 
ton;  soluble  silver,  16.33  oz.  per  ton,  or  54.9  per  cent. 

Eighth  hour.  —  The  same  moderate  roasting  temperature. 
No  more  sulphurous  acid  gas  could  be  noticed.  The  ore  com- 
menced to  look  dead.  The  color  of  the  sample  when  cold  did  not 
change  in  this  nor  in  the  following  hours. 

Assay  of  ore,  30.62  oz.  per  ton;  after  leaching,  13.55  oz.  per 
ton;  soluble  silver,  17.05  oz.  per  ton,  or  55.7  per  cent. 

Ninth  hour.  —  The  temperature  was  slightly  increased,  but 
the  ore  did  not  fume,  nor  could  sulphurous  acid  be  noticed. 

Wash-water:  a  very  heavy,  yellowish-white  precipitate. 

Hypo-solution  (1  per  cent.):  a  heavy  precipitate,  consisting 
principally  of  lead,  zinc  and  silver,  no  copper. 

Assay  of  ore,  29.30  oz.  per  ton;  after  leaching,  12.09  oz.  per 
ton;  soluble  silver,  17.21  oz.  per  ton,  or  58.8  per  cent. 

Tenth  hour.  —  Temperature  again  slightly  increased;  no  fumes, 
no  sulphurous  acid  noticeable.  The  ore  remains  fine  and  loose; 
no  lumps. 

Assay  of  ore,  29.16  oz.  per  ton;  after  leaching,  12.24  oz.  per 
ton;  soluble  silver,  16.92  oz.  per  ton,  or  58.0  per  cent. 


124  HYDROMETALLURGY   OF  SILVER 

Eleventh  hour.  —  During  this  hour  the  temperature  was  con- 
siderably increased,  bright  red;  the  ore  remained  fine  and  loose; 
no  lumps;  no  sulphurous  acid  noticeable. 

Wash-water;  a  yellowish-white  precipitate  of  very  clear  color. 
Assay  of  ore,  30.03  oz.  per  ton;  after  leaching,  13.26  oz.  per 
ton;  soluble  silver,  16.77  oz.  per  ton,  or  55.8  per  cent. 

As  the  ore  during  this  hour  presented  the  appearance  of  a 
dead  roasted  ore,  oxidation  was  not  carried  any  further,  but  6 
per  cent,  of  salt  was  added  at  the  end  of  the  eleventh  hour,  to 
make  some  further  observations. 

At  the  end  of  the  eleventh  hour  and  just  after  adding  6  per  cent, 
salt.  —  The  salt  was  added,  quickly  stirred  and  a  sample  taken. 

Wash-water:  dark -colored  precipitate. 

Assay  of  ore,  28.74  +  6  per  cent.  =  30.46  oz.  per  ton;  after  leach- 
ing, 11.66  +  6  per  cent.  =  12.36  oz.  per  ton;  soluble  silver,  18.80 
oz.,  or  59.4  per  cent. 

Twelfth  hour.  —  The  temperature  lowered  for  chloridizing. 
The  ore  commenced  to  fume  after  the  salt  was  added,  but  the 
fumes  were  very  thin  and  light. 

Wash-water:  black  precipitate. 

Assay  of  ore,  28.13  +  6  per  cent.  =  29.81  oz.  per  ton;  after 
leaching,  13.12  +  6  per  cent.  =  13.90  oz.  per  ton;  soluble  silver, 
15.90  oz.  per  ton,  or  53.3  per  cent. 

Thirteenth  hour.  —  Same  temperature;  a  mild,  faint  odor  of 
chlorine  perceptible;  no  lumps,  and  when  the  ore  was  left  in  a 
heap  to  cool  it  did  not  harden  like  other  charges  roasted  with 
salt.  At  the  end  of  this  hour  the  charge  was  removed  from  the 
furnace. 

Wash-water:  dark  precipitate,  almost  black. 

Assay  of  ore,  28.42  +  6  per  cent.  =  30.12  oz.  per  ton;  after 
leaching,  11.07  +  6  per  cent.  =  11. 73  oz.  per  ton;  soluble  silver, 
18.39  oz.,  or  61  per  cent. 

I  found  that  the  assay  value  of  the  ore  during  oxidizing 
roasting  dropped  from  35.57  oz.  per  ton  to  29.74  oz.  This  was 
caused  principally  by  an  increase  in  weight  which  the  ore  sus- 
tained by  oxidation,  especially  the  lead  and  zinc  of  which  a  large 
percentage  is  converted  into  sulphate  if  roasted  at  such  a  low 
heat.  This  took  place  principally  during  the  seventh  hour, 
when  the  ore  assumed  a  dark  red-brown  color.  I  further  found 
that  during  oxidizing  roasting  considerable  of  the  silver  is  con- 


CHLORIDIZING  OF  ARGENTIFEROUS  ZINC-LEAD  ORE     125 

verted  into  a  state  in  which  it  is  soluble  in  sodium  hyposulphite. 
It  shows  first  at  the  end  of  the  third  hour  (3.9  per  cent.)  and  gradu- 
ally increases  until  it  reaches  its  maximum  (58.8  per  cent.)  at  the 
end  of  the  ninth  hour,  and  then  diminishes  during  the  next  two 
hours  to  55.8  per  cent. 

By  adding  6  per  cent,  of  salt  and  continuing  to  roast  for  two 
hours,  the  percentage  of  soluble  silver  increased  only  2.2  per 
cent,  above  the  maximum  obtained  in  oxidizing  roasting,  not- 
withstanding that  at  the  time  the  salt  was  added  the  ore  was  not 
dead  roasted,  but  still  contained  sulphates  soluble  in  water. 
The  salt  soluble  in  water,  however,  was  zinc  sulphate,  which 
does  not  act  on  the  salt;  the  iron  sulphate  was  decomposed  by 
that  time.  This  experiment  illustrates  the  great  importance  of 
adding  the  salt  at  a  certain  time  during  oxidizing  roasting  — 
that  is,  at  a  time  before  the  iron  sulphate  is  decomposed  and  the 
oxidation  of  the  lead  and  zinc  sulphides  has  fairly  advanced. 

Only  a  part  of  the  soluble  silver  in  the  oxidized  ore  was  present 
as  a  sulphate.  From  a  sample  of  nine  hours'  oxidizing,  when 
leached  with  water,  only  7.39  oz.  silver  per  ton  could  be  ex- 
tracted, while  the  sodium  hyposulphite  extracted  17.21  oz.  per  ton. 
In  other  words,  of  the  58.8  per  cent,  of  soluble  silver  only  25.2 
per  cent,  was  sulphate  of  silver,  while  the  balance  of  33.6  per 
cent,  was  some  other  silver  salt,  not  soluble  in  water  but  soluble 
in  sodium  hyposulphite,  probably  silver  antimonate,  though  the 
raw  ore  contained  only  one-half  per  cent,  of  antimony. 

TREATING  THE  OXIDIZED  ORE  WITH  CUPRIC  CHLORIDE 

A  sample  of  the  nine  hours'  oxidizing  roasting  was  moistened 
with  a  dilute  solution  of  cupric  chloride  and  left  in  that  condition 
for  three  hours,  then  leached  with  water  and  hypo  solution: 
Assay  of  ore,  29.30  oz.  per  ton;  after  leaching  7.29  oz.  per  ton; 
soluble  silver,  22.01  oz.  or  75.2  per  cent.  Increased  extraction 
by  using  cupric  chloride  4.8  oz.  per  ton,  or  19.4  per  cent. 

CONSUMPTION    OF    WOOD    IN    THE    REVERBERATORY    FURNACE 

The  furnaces  were  built  in  pairs,  two  being  connected  with 
one  flue.  During  four  weeks  the  wood  consumed  by  one  pair  of 
these  furnaces  was  weighed.  During  this  time  507  tons  were 
roasted  at  a  consumption  of  672  cargas  of  wood,  or  1.3  cargas  per 


126  HYDROMETALLURGY  OF  SILVER 

ton.     If  expressed  in  cords,  we  find  that  one  cord  of  wood  roasted 
nine  tons  of  ore,  which  is  an  exceedingly  small  consumption. 

COST  OF  ROASTING  IN  THE  REVERBERATORY  FURNACE 

Statement  of  cost  for  two  two-story  furnaces  roasting  18  tons 
per  day: 

Labor $22.60 

23.4  cargas  of  wood  at  75^ 17.55 

4  per  cent,  salt,  1440  Ib.  at  1.27^ , 18.28 

Tools,  etc 4.00 

Management,  office,  mechanics,  etc 3.77 

$66\20-18=$3.67 
Cost  per  ton $3.67  Mexican  currency. 

CONCLUSIONS 

To  form  an  opinion  as  to  which  roasting  furnace  is  the  most 
suitable  for  the  San  Francisco  del  Oro  ore,  we  have  to  take  into 
consideration  only  the  modified  Howell  and  the  reverberatory. 
The  Stetefeldt  did  not  roast  the  ore,  and  the  Bruckner  was  not 
tried,  because  I  knew  by  experience  that  a  large  Bruckner  furnace 
could  not  roast  more  than  five  or  six  tons  per  day  of  such  a 
heavy  ore,  and  therefore  refrained  from  incurring  the  expense  of 
erecting  a  Bruckner  furnace  just  for  experimental  purposes. 

Both  furnaces,  the  modified  Howell  and  the  reverberatory, 
gave  about  the  same  results,  and  the  loss  by  volatilization  was 
also  nearly  the  same.  The  cost  of  roasting,  however,  is  different. 
The  cost  per  ton  of  ore  in  the  modified  Howell  is  found  to  be  $4.50, 
and  in  the  reverberatory  $3.67.  Difference  in  favor  of  reverbera- 
tory, 83c.  per  ton.  Besides  this,  the  reverberatory  furnace 
creates  much  less  flue-dust  than  the  Howell.  The  latter  forms  a 
great  deal  of  flue-dust  which  is  far  from  being  roasted,  even  if 
provided  with  an  auxiliary  fireplace.  The  labor  question  has 
also  to  be  taken  into  consideration.  The  reverberatory  requires 
more  manual  labor,  and  in  a  locality  where  labor  is  scarce,  as  is 
often  the  case  in  Mexico,  it  may  be  more  advantageous,  under 
certain  conditions,  to  adopt  the  Howell  notwithstanding  the 
greater  cheapness  of  the  reverberatory. 


XII 

CHLORIDIZING  OF  CALCAREOUS  ORES 

I  WAS  engaged  by  the  Anglo-Mexican  Mining  Company  to 
investigate  and  improve  the  roasting  of  their  ores  at  Yedras, 
Sinaloa,  Mexico. 

The  ore  is  treated  by  the  lixiviation  process  with  sodium 
hyposulphite,  and  consists  of  argentiferous  arsenical  pyrites, 
fine-grained  black  zinc  blende,  arsenical  fahlerz,  some  iron  py- 
rites, and  occasionally  ruby  silver,  while  the  gangue  consists  of 
silicious  limestone  and  calcspar. 

The  chloridizing  roasting  of  highly  calcareous  silver  ores  in 
combination  with  argentiferous  arsenical  pyrites  has,  like  the  ore 
treated  in  the  preceding  chapter,  not  often,  if  ever,  been  the 
subject  of  a  thorough  investigation  on  a  large  scale,  and  in  the 
following  pages  the  results  and  observation  of  such  an  investi- 
gation are  given  in  detail. 

The  only  analysis  of  the  ore  which  I  could  obtain  is  the  fol- 
lowing, made  in  San  Francisco  several  years  before  the  experi- 
ments were  made: 

Per  cent. 

Silica 15.13 

Sulphur 13.31 

Arsenic 9.82 

Iron 17.33 

Alumina 1 .35 

Zinc 4.92 

Lead 1.78 

Carbonate  of  lime 33.78 

Magnesia 2.58 

This  analysis,  however,  does  not  represent  the  average  of  the 
ore  which  was  delivered  from  the  mine  to  the  mill.  Frequent 
concentration  tests  showed  that  the  mill  ore  contained  much 
more  gangue  matters  than  the  analysis  shows.  However,  it  may 
serve  to  give  a  general  idea  of  the  ore.  While  experimenting  I 

127 


128  HYDROMETALLURGY   OF  SILVER 

felt  very  much  the  want  of  a  chemical  laboratory  and  a  chemist 
at  the  works. 

The  roasting  facilities  at  the  Yedras  mill  consisted  of  four 
revolving  Bruckner  cylinder  furnaces,  each  16  ft.  long,  and  eight 
long  reverberatory  furnaces.  The  Briickners  had  been  abandoned 
for  several  years,  because  former  operators,  so  I  was  told,  could 
not  get  satisfactory  res'ults  with  them.  The  chlorination  was 
exceedingly  low  (40  to  50  per  cent.),  and  the  loss  of  silver  by 
volatilization  extremely  high,  while  the  main  part  of  the  roasted 
ore  was  rolled  up  into  balls  ranging  from  the  size  of  an  orange 
up  to  15  in.  and  upward  in  diameter,  the  inside  of  which  was 
not  roasted.  The  subsequently  erected  reverberatory  furnaces 
also  gave  very  poor  results  —  65  to  70  per  cent,  chlorination,  with 
a  loss  of  silver  by  volatilization  of  20  to  35  per  cent,  though 
fewer  lumps  were  formed. 

ROASTING  IN  THE  BRUCKNER  FURNACES 

I  knew  by  experience  that  as  a  rule  ores  can  be  roasted  with 
less  loss  of  silver  by  volatilization  in  the  Bruckner  than  in  the 
reverberatory  furnace,  and  as  the  enormous  loss  of  silver  which 
the  ore  had  sustained  in  the  reverberatory  was  the  most  important 
question,  I  started  the  long-abandoned  Bruckner  furnaces  again. 
The  previous  failures  with  them  I  ascribed  to  the  application  of 
too  high  a  temperature  and  an  insufficient  supply  of  air.  The 
formation  of  balls  I  expected  to  diminish  by  judiciously  regulating 
the  heat  and  the  revolving  speed  of  the  furnace,  and  if  I  succeeded 
in  chloridizing  well  these  lumps,  their  formation  would  not  be  a 
serious  matter,  for  they  could  be  pulverized  in  a  ball-mill  before 
charging  the  ore  into  the  leaching  vats. 

My  improvements  on  the  Bruckner  cylinder,  which  consist  of 
a  fireplace  and  flue  arrangement  attached  to  each  end  of  the 
furnace,  enable  me  to  apply  the  flame  alternately  through  either 
end,  and  thus  to  use  much  longer  cylinders,  and  provide  fully  for 
the  free  access  of  air  between  the  fireplace  and  the  throat  of  the 
furnace,  which,  strange  to  say,  is  not  the  case  with  the  common 
Bruckner  furnaces.  In  a  reverberatory  furnace  air  can  enter 
through  the  working  doors;  Stetefeldt  provided  his  furnace  with 
air-channels  through  which  the  required  supply  can  be  regulated; 
the  Howell  furnace  is  provided  with  an  air-door;  but  the  common 


CHLORIDIZING  OF  CALCAREOUS   ORES  129 

Bruckner  furnace  has  no  proper  means  of  regulating  the  supply 
of  air,  and  either  the  fire-door  must  remain  open  or  only  a  very 
limited  amount  of  air  can  enter  the  furnace.  If  the  whole  fire-box 
is  built  of  brick  it  is  easy  enough  to  make  the  proper  change,  but  if 
it  is  made  of  boiler-iron  lined  with  brick,  it  is  not  easy  to  make  the 
necessary  alterations  in  a  remote  mining  camp.  In  this  case,  the 
fire-boxes  being  of  brick,  I  added  12  in.  to  their  length,  moved 
the  grate  bars  toward  the  front,  and  inserted  an  air-channel 
behind  the  new  fire-bridge.  The  fresh  air  entering  through  this 
chamber  not  only  assisted  in  oxidizing  the  ore,  but  also  aided 
the  combustion  of  the  fuel.  The  speed  of  the  furnace,  which  had 
been  2J  revolutions  per  minute,  was  reduced  to  one  revolution 
in  If  minutes,  which  materially  reduced  or  altogether  avoided  the 
formation  of  large  balls  and  diminished  the  quantity  of  dust, 
while  on  the  other  hand  the  speed  was  sufficient  to  expose  every 
particle  of  the  ore  to  the  action  of  air  and  heat;  in  fact,  a  still 
slower  speed,  possibly  even  one  revolution  in  three  or  four  minutes, 
would  have  been  preferable  had  it  been  obtainable. 

The  ore  was  crushed  dry  in  the  stamp  battery  with  7  per  cent, 
of  salt,  and  passed  through  a  24-mesh  screen.  The  furnace  charge 
was  4  to  4J  tons,  and  a  strong  fire  was  kept  in  order  to  quickly 
ignite  the  ore.  In  about  an  hour,  and  before  the  sulphur  com- 
menced to  burn,  heavy  arsenic  fumes  were  given  off,  entirely 
obscuring  the  interior  of  the  furnace.  After  one  and  one-half  or 
two  hours  a  sulphur  flame  could  be  observed  to  enter  the  flue, 
which  was  a  sign  that  the  sulphurets  were  sufficiently  ignited  to 
continue  combustion  without  the  aid  of  fire.  The  fire  was  then 
allowed  to  go  out  and  the  fire  and  air-doors  were  kept  wide  open. 
The  temperature  gradually  increased  by  the  combustion  of  the 
sulphur  until  it  reached  a  certain  maximum,  at  which  it  remained 
for  several  hours.  After  two  or  three  hours  the  arsenic  period  was 
over,  the  heavy  fumes  disappeared,  the  interior  of  the  furnace 
became  clear,  and  the  glow  of  the  ore  presented  a  beautiful  pink 
color.  No  chemical  loss  of  silver  took  place  during  the  arsenic 
period.  The  oxidizing  period  continued  for  three  to  four  hours 
more,  the  ore  remaining  at  about  the  same  temperature  through- 
out, then  fumes  commenced  to  rise  from  the  ore,  and  gradually 
increased,  but  never  became  so  dense  as  during  the  arsenic  period, 
though  still  dense  enough  to  make  the  interior  of  the  furnace 
invisible.  The  chloridizing  period  had  commenced,  and  the  odor 


130  HYDROMETALLURGY  OF  SILVER 

of  sulphurous  acid  and  chlorine  could  be  observed  in  the  samples 
taken  during  the  first  period;  but  later  the  odor  of  sulphurous 
acid  disappeared  and  only  chlorine  could  be  detected.  In  many 
instances,  however,  no  smell  of  chlorine  could  be  noticed  during 
the  whole  time  of  roasting,  as  explained  further  on.  Looking 
through  the  open  fireplace,  which  by  this  time  had  cooled  down, 
the  whole  interior  seemed  to  be  glowing,  though  the  ore  itself  was 
not  visible  on  account  of  the  zinc  fumes,  which  were  illuminated 
by  the  glow  of  the  red-hot  ore  and  made  a  beautiful  sight,  and 
at  the  same  time  afforded  a  good  opportunity  to  observe  the 
temperature  in  the  furnace.  So  soon  as  a  decrease  in  temperature 
was  noticed,  the  fire  was  started  again  and  kept  up  for  three  to 
four  hours,  when  the  charge  was  considered  finished.  It  took 
altogether  from  twelve  to  fourteen  hours  to  roast  a  charge. 

During  the  oxidizing  period  the  ore  maintained  an  almost 
level  position  in  the  cylinder  and  had  a  liquid-like  appearance. 
The  ore  particles  on  the  surface,  however,  could  be  seen  constantly 
moving;  on  the  side  where  the  furnace  moves  up  fresh  ore  came 
to  the  surface  as  if  emerging  from  a  liquid,  moved  slowly  across, 
and  sank  as  soon  as  it  touched  the  down-moving  side  of  the 
cylinder.  The  ore  increased  considerably  in  volume  during  the 
first  part  of  the  chloridizing  period;  but  after  it  had  reached 
the  maximum  it  commenced  to  shrink  again,  assuming  a  heavy 
sandy  condition,  where  before  it  was  loose  and  woolly,  and  finally 
occupied  no  more  space  in  the  furnace  than  the  raw  ore  did. 
A  short  time  before  the  second  fumes  commenced  to  rise  the 
charge  assumed  a  more  inclined  position,  attaining  nearly  45  deg. 

There  was  a  great  and  puzzling  irregularity  in  the  results. 
Sometimes  a  number  of  successive  charges  gave  satisfactory 
results;  then  at  once  the  chlorination  dropped  without  any 
apparent  cause,  the  ore,  the  amount  of  salt,  the  temperature  and 
the  treatment  being  unchanged.  The  only  noticeable  difference 
was  the  amount  of  silver  contained  in  the  ore;  and  as  a  rule  the 
richer  ore  gave  the  better  results,  though  it  was  evident  the 
quantity  of  silver  as  such  could  not  influence  the  result  materially. 
Close  investigation  finally  showed  the  true  cause  of  the  trouble 
to  be  the  more  or  less  favorable  proportion  between  the  gangue 
(carbonate  of  lime)  and  the  sulphureted  matters  in  the  ore;  and 
this  explanation  corresponded  with  the  fact  noticed  that  richer 
ores  usually  gave  better  results  than  poorer  ores.  No  marked 


CHLORIDIZING  OF  CALCAREOUS  ORES 


131 


difference  between  the  two  could  be  noticed  in  an  inspection  of 
the  ore  before  it  went  into  the  battery. 

The  following  tables  of  the  results  obtained  with  ores  con- 
taining less  and  ores  containing  more  carbonate  of  lime  illustrate 
these  differences: 


ORE  CONTAINING  LESS  CARBONATE  OF  LIME 

(7  per  cent,  salt  mixed  in  the  battery.) 


No.  OF 
CHARGE 

VALUE  OF  RAW 
ORE  PER  TON 

VALUE  OF  ROASTED 
ORE  PER  TON 

VALUE  OF  LEACH 
TAILINGS  PER  TON 

CHLORINATION 

Oz.  Silver 

Oz.  Silver 

Oz.  Silver 

Per  Cent. 

65 

72.12 

76.12 

16.20 

79.6 

66 

(a) 

64.86 

12.60 

80.5 

67 

65.64 

64.80 

9.00 

86.2 

68 

68.04 

68.46 

10.56 

84.6 

69 

66.30 

72.16 

13.68 

81.1 

70 

66.78 

73.62 

12.36 

86.0 

71 

68.16 

71.16 

11.88 

83.3 

72 

66.24 

68.64 

12.90 

81.3 

73 

68.52 

73.82 

14.70 

80.0 

74 

66.00 

75.24 

11.16 

85.4 

75 

64.38 

66.42 

12.06 

81.9 

76 

63.78 

66.30 

9.86 

85.9 

77 

68.28 

68.10 

11.42 

83.3 

78 

58.50 

61.02 

10.80 

82.4 

862.74 

970.72 

169.18 

1161.5 

(a)  Sample  of  raw  ore  lost. 

Average  of  raw  ore 66.36  oz.  per  ton. 

Average  of  roasted  ore 69.33  oz.  per  ton. 

Average  of  leach  tailings 12.08  oz.  per  ton. 

Average  of  chlorination 82.96  per  cent. 

The  roasted  ore  contained  2.97  oz.  more  silver  per  ton  than 
the  raw  ore. 


132 


HYDROMETALLURGY  OF  SILVER 


ORE  CONTAINING  MORE  CARBONATE  OF  LIME 

(7  per  cent,  salt  mixed  in  the  battery.) 


No.  OF 
CHARGE 

VALUE  OF  RAW 
ORE  PER  TON 

VALUE  OF  ROASTED 
ORE  PER  TON 

VALUE  OF  LEACH 
TAILINGS  PER  TON 

CHLORINATION 

Oz.  Silver 

Oz.  Silver 

Oz.  Silver 

Per  Cent. 

79 

61.06 

60.36 

15.90 

73.7 

80 

62.46 

64.20 

19.14 

71.2 

81 

60.18 

64.08 

16.14 

74.9 

82 

56.70 

56.52 

11.76 

79.2 

83 

55.08 

55.44 

12.72 

77.1 

84 

57.12 

53.88 

20.80 

61.4 

85 

60.34 

58.98 

14.52 

75.4 

86 

59.64 

57.60 

20.04 

65.3 

87 

58.50 

56.10 

18.18 

67.6 

88 

49.92 

52.86 

15.18 

71.3 

89 

55.38 

54.90 

12.84 

76.7 

90 

55.80 

52.56 

12.90 

76.2 

91 

57.60 

51.36 

9.12 

81.3 

92 

57.00 

55.80 

13.20 

76.4 

93 

55.02 

55.74 

15.00 

73.1 

94 

58.32 

56.64 

14.64 

74.2 

95 

57.00 

57.36 

7.62 

86.8 

977.12 

964.38 

249.70 

1261.8 

Average  of  raw  ore 57.47  oz.  per  ton. 

Average  of  roasted  ore 56.72  oz.  per  ton. 

Average  of  leach  tailings 14.68  oz.  per  ton. 

Average  of  chlorination 74.22  per  cent. 

The  raw  ore  contained  0.75  oz.  more  silver  per  ton  than  the 
roasted  ore. 

Comparing  the  average  results,  they  are  found  to  be  decidedly 
in  favor  of  the  higher  sulphureted  ore.  To  obtain  further  infor- 
mation I  concentrated  some  of  the  ore,  and  made  mixtures  of 
certain  proportions  of  concentrates  and  barren  gangue,  mostly 
carbonate  of  lime,  and  roasted  these  different  mixtures  as  well  as 
the  concentrates  in  the  muffle,  treating  all  samples  alike,  using 
7  per  cent,  of  salt,  and  roasting  each  half  an  hour.  The  concen- 
trates used  were  obtained  from  ore  which  did  not  represent  the 
average  richness,  assaying  only  35.70  oz.  per  ton,  and  therefore 
the  different  proportions  I  made  were  much  poorer  in  silver  than 
I  afterward  found  corresponding  ones  in  the  bulk  of  the  run. 
Only  the  concentrates  were  assayed,  the  values  of  the  mixtures 
being  calculated. 

These  tests  showed  that  the  silver-bearing  minerals  of  this 
ore  do  not  offer  any  difficulties  to  a  good  chloridizing  roasting; 


CHLORIDIZING  OF  CALCAREOUS  ORES 


133 


on  the  contrary,  they  are  easy  to  roast  to  a  high  percentage 
without  showing  any  tendency  to  ball.  The  difficulties  the  ore 
offered  were  therefore  caused  by  the  gangue;  and  the  great  varia- 
tion in  the  results  was  due  to  a  greater  or  less  favorable  proportion 
of  sulphureted  matters  to  lime. 


No.  OF 
SAMPLE 

CONCENTRATES 

BARREN 
GANGUE 

VALUE  OF 
MIXTURE 
PER  TON 

VALUE  OF 
LEACH  TAILS 
PER  TON 

CHLORINA- 

TION 

Per  Cent. 

Per  Cent. 

Oz.  Silver 

Oz.  Silver 

Per  Cent. 

1  (a) 

100.0 

96.0 

2.91 

97.0 

2(6) 

75.0 

25.0 

72.0 

5.38 

92.6 

3(c) 

62.5 

37.5 

60.0 

4.72 

92.2 

4(d) 

50.0 

50.0 

48.0 

5.38 

88.8 

5(e) 

25.0 

75.0 

24.0 

5.47 

77.2 

(a)  Fuming  profusely  during  chloridizing  period;  strong  and  pure  smell  of 
chlorine;  a  great  deal  of  base-metal  chlorides  soluble  in  water  was  formed; 
no  tendency  to  form  lumps  shown.  (6)  During  chloridizing  heavy  fumes; 
distinct  but  moderately  strong  smell  of  chlorine;  no  base-metal  chlorides 
formed;  no  tendency  to  form  lumps  shown,  (c)  Much  less  fumes;  very 
little  chlorine  observable ;  none  in  water-soluble  base-metal  chlorides ;  no 
tendency  to  form  lumps,  (d)  Very  little  fumes;  no  chlorine;  no  base-metal 
chlorides;  showed  tendency  to  form  lumps,  (e]  No  fumes  except  at  a  high 
heat,  and  then  but  little;  no  chlorine ;  none  in  water-soluble  chlorides; 
showed  much  tendency  to  form  lumps. 

Carbonate  of  lime  in  presence  of  heated  sulphureted  minerals 
will  change  partly  into  calcium  sulphate,  which  does  not  act  on 
sodium  chloride,  and  partly  into  caustic  lime,  which  decomposes 
the  metal  sulphates  and  chlorides,  and  also,  though  less  rapidly, 
the  silver  chloride.  If,  however,  carbonate  of  lime  is  greatly  in 
excess,  only  a  very  small  amount,  if  any,  of  base-metal  sulphates 
is  formed  to  decompose  sodium  chloride,  the  greater  part  being 
changed  directly  into  oxides.  If  no  salt  is  present  these  sulphates 
are,  of  course,  quickly  changed  into  oxides.  For  this  reason  no 
chlorine  can  be  detected  if  the  salt  is  added  after  the  ore  has  been 
subjected  to  a  partial  oxidizing  roasting. 

Some  of  the  more  calcareous  charges  were  followed  closely, 
but  at  no  stage  of  the  roasting  could  any  soluble  salts  or  chlorine 
be  detected  if  the  salt  was  added  after  the  ore  had  been  oxidizing 
for  some  time.  If  the  salt  was  pulverized  with  the  ore  in  the 
battery  traces  of  such  salts  were  found,  but  never  in  large  quanti- 
ties. 

As  there  were  neither  sulphates,  which  at  a  high  temperature 


134  HYDROMETALLURGY  OF  SILVER 

decompose  salt,  nor  chlorides,  which  at  a  high  heat  give  off 
chlorine,  an  increased  temperature  during  the  last  part  of  the 
roasting  could  not  be  of  any  benefit.  In  fact,  it  had  even  a  very 
bad  effect.  The  caustic  lime,  which  at  a  low  temperature  seemed 
to  be  comparatively  indifferent  to  silver  chloride,  decomposed  it 
energetically  at  a  high  temperature.  The  roasting  of  such  ores 
had  therefore  to  be  finished  at  as  low  a  temperature  as  possible, 
contrary  to  the  usual  practice.  This  fact  must  be  taken  into 
consideration  in  constructing  long  reverberatory  furnaces.  The 
arch  of  the  first  and  second  hearths  nearest  to  the  fire  should  be 
very  high.  (Figs.  4,  5  and  6.) 

Charge  No.  86  of  the  more  calcareous  ore  was  roasted  for 
eight  hours  without  fire;  then,  when  the  ore  commenced  to  lose 
heat,  a  second  fire  was  applied  for  three  hours.  Before  starting 
the  second  fire  a  sample  was  taken  through  the  entire  length  of 
the  furnace,  and  the  chlorination  was  found  to  be  73.4  per  cent., 
and  after  three  hours'  second  fire  it  was  only  65.3  per  cent. 

The  moderately  increased  temperature  at  that  stage  of  the 
roasting  reduced  the  chlorination  8  per  cent.  A  number  of  such 
observations  were  made  and  finally  the  mode  of  roasting  was 
changed,  and  the  ore  was  allowed  to  roast  entirely  by  the  oxidation 
of  the  sulphureted  matters  without  the  application  of  a  second 
fire. 

Highly  calcareous  charges,  containing  only  about  20  to  30 
per  cent.  sulphurets,had  a  dead  and  sandy  appearance,  no  chlorine 
could  be  noticed,  and  it  required  a  high  heat  to  make  the  ore 
fume.  These  fumes  were  light  and  were  probably  zinc  oxide. 
Not  even  at  a  very  high  heat  could  any  chlorine  be  detected. 
If,  however,  a  very  strong  draft  of  air  was  allowed  to  pass  through 
the  furnace  from  the  beginning  and  during  the  whole  time  of 
roasting,  some  chlorine  was  generated  and  could  be  detected. 
The  ore  fumed  moderately  at  a  low  temperature  and  70  to  75 
per  cent,  of  the  silver  was  rendered  soluble  in  sodium  hyposulphite. 
This  indicated  that  by  a  rapid  supply  of  air  more  sulphuric  and 
less  sulphurous  acid  was  generated,  and  that  some  of  the  sodium 
chloride  was  decomposed  by  the  former. 

The  beneficial  effect  of  a  great  surplus  of  air  was  still  more 
noticeable  with  highly  sulphureted  ore  in  the  reverberatory  furnace. 
Differences  as  high  as  25  to  30  per  cent,  in  chlorination  were  ob- 
served between  two  charges  of  the  same  ore,  which  were  treated 


CHLORIDIZING  OF  CALCAREOUS  ORES  135 

alike,  except  that  one  furnace  had  excessive  draft  while  the 
other  had  insufficient  air.  Silicious  ores  can  be  well  chloridized 
with  a  moderate  supply  of  air,  while  calcareous  ores  need  an  excess. 
This  fact  is  of  great  importance,  and  due  attention  should  be  paid 
to  it  in  constructing  furnaces. 

An  ore  which  by  hand  concentration  showed  40  to  45  per  cent, 
sulphurets  generated  considerable  chlorine;  it  fumed  more  freely 
and  at  a  lower  temperature;  and  though  no  base-metal  chlorides 
were  formed,  80  to  86  per  cent,  of  the  silver  was  chloridized  when 
temperature,  time,  and  supply  of  air  were  properly  regulated. 
Ore  showing  by  hand  concentration  50  to  55  per  cent,  of  sulphurets 
permitted  a  chlorination  of  90  per  cent,  and  over. 

In  the  Bruckner  furnace  the  ore  behaved  very  differently, 
according  as  the  salt  was  added  in  the  battery  or  in  the  furnace, 
and  it  was  therefore  necessary  to  examine  each  case  separately. 

ADDING  THE  SALT  IN  THE  FURNACE 

If  the  ore  was  allowed  to  oxidize  without  salt  a  certain  per- 
centage of  the  silver  would  be  rendered  soluble  in  sodium  hypo- 
sulphite, but  not  in  water,  owing  to  the  arsenic  contained  in  the 
ore.  This  change  takes  place  principally  at  the  very  beginning 
of  roasting  and  at  a  very  low  temperature.  A  sample  taken 
from  charge  No.  128  an  hour  after  the  combustion  of  the  sul- 
phurets had  commenced,  while  the  arsenic  was  still  fuming 
strongly  and  oxidization  had  but  just  commenced,  the  appear- 
ance of  the  sample  being  still  that  of  raw  ore,  yielded  no  Jess  than 
44.58  per  cent,  of  the  silver  with  sodium  hyposulphite.  Con- 
tinuing the  oxidation,  the  amount  of  soluble  silver  increased, 
but  not  so  rapidly  as  in  the  first  hour,  as  was  shown  by  the  follow- 
ing tests: 

Charge  No.  128,  after  1  hour  oxidizing  roasting,  yielded  44.58 
per  cent,  of  its  silver  with  sodium  hyposulphite. 

Charge  No.  128,  after  1J  hours  oxidizing  roasting,  yielded 
43.44  per  cent,  of  its  silver  with  sodium  hyposulphite. 

Charge  No.  128,  after  2J  hours  oxidizing  roasting,  yielded 
45.32  per  cent,  of  its  silver  with  sodium  hyposulphite. 

Charge  No.  128,  after  3£  hours  oxidizing  roasting,  yielded 
57.66  per  cent,  of  its  silver  with  sodium  hyposulphite. 

No  marked  increase  in  soluble  silver  could  be  noticed  by  con- 
tinuing the  oxidation  still  further. 

A  certain  percentage  of  silver  can  be  extracted  from  some 


136  HYDROMETALLURGY  OF  SILVER 

ores  with  sodium  hyposulphite  without  roasting,  and  in  order  to 
determine  whether  this  extractable  form  of  silver  was  originally 
contained  in  the  ore,  or  was  formed  during  even  such  a  slight 
oxidizing  roasting  as  charge  No.  128  showed,  where  in  one  hour 
44.58  per  cent,  became  extractable,  a  sample  of  raw  ore  was 
lixiviated,  but  not  even  a  fraction  of  1  per  cent,  of  silver  could  be 
extracted.  As  this  silver  combination  soluble  in  sodium  hypo- 
sulphite was  principally  formed  in  the  first  stage  of  oxidizing 
roasting,  during  the  time  the  arsenic  was  being  expelled,  it  was 
undoubtedly  silver  arsenate. 

When  the  salt  was  added  at  any  time  during  the  oxidizing 
period  and  well  stirred  in  with  a  hoe,  the  percentage  of  soluble 
silver  either  increased  suddenly  a  few  per  cent,  or  decreased,  but 
in  neither  case  could  this  percentage  be  maintained  or  further 
increased.  Decomposition  commenced  immediately  and  the  per- 
centage of  extractable  silver  diminished  rapidly,  when,  after  two 
to  four  hours,  according  to  the  amount  of  salt  added,  the  maxi- 
mum was  reached. 

In  the  following  tables  are  given  the  detailed  history  of  three 
charges,  roasted  with  10,  7,  and  4  per  cent,  of  salt  respectively. 
The  maximum  of  decomposition  was  reached  in  two,  three,  and 
four  hours: 


CHLORIDIZING  OF  CALCAREOUS  ORES 


137 


CHARGE  NO.  116 
(10  per  cent,  salt  added  during  roasting.) 


DESCRIPTION  (a) 

VALUE  OF 
ROASTED  ORE 
CONTAINING 
SALT 
PER  TON 

VALUE  OF 
LEACH  TAIL- 
INGS 
PER  TON 

SOLUBLE 
SILVER 

Raw  ore  Oz.  per  ton,  62.52 

Oz.  Silver 

Oz.  Silver 

Per  Cent. 

After  5  hours  and  20  minutes  roast- 
ing and  just  before  adding  salt 
Oz  per  ton  61  32 

The    same    calculated   to   contain 
10  per  cent  salt.  .Oz.per  ton,  55.19 
Directly    after    adding    salt    and 
stirred  

5724 

28.80 
26  04 

53.1 
54  5 

Twenty  minutes  after  adding  salt  .  .  . 
One  hour  after  adding  salt  
Two  hours  after  adding  salt  

62.04 
61.98 
63  24 

31.92 
39.48 
46,68 

48.6 
36.3 
26.2 

Three  hours  after  adding  salt  
Four  hours  after  adding  salt  
Five  hours  after  adding  salt  

62.10 
64.32 
64.20 

42.90 
41.64 
40.02 

31.0 
35.3 
37.7 

(a)  After  5  hours  and  20  minutes  oxidizing  roasting  10  per  cent,  salt  was 
added  and  the  charge  stirred  with  hoes.  The  maximum  of  decomposition 
was  reached  in  2  hours  after  the  salt  was  added,  at  which  time  51.92  per 
cent,  of  the  soluble  silver  was  rendered  insoluble.  A  decomposition  of  10.82 
per  cent,  took  place  in  the  first  20  minutes;  in  the  next  40  minutes,  22.57 
per  cent.;  in  the  next  60  minutes,  18.53  per  cent. 


138 


HYDROMETALLURGY  OF  SILVER 


CHARGE  NO.  123 
(7  per  cent,  salt  added  during  roasting.) 


DESCRIPTION  (a) 

VALUE  OF 
ROASTED  ORE 
CONTAINING 
SALT 
PER  TON 

VALUE  OF 
LEACH  TAIL- 
INGS 
PER  TON 

SOLUBLE 
SILVER 

Raw  ore  Oz.  per  ton,  58.62 

Oz.  Silver 

Oz.  Silver 

Per  Cent. 

After  4  hours  roasting  and  just  be- 
fore adding  salt  .  .  Oz.  per  ton,  59.70 

21  36 

64  27 

One  hour  after  adding  salt 

59  84 

27  00 

*4  7s 

Two  hours  after  adding  salt  
Three  hours  after  adding  salt  

56.04 
55.92 

35.40 
41  16 

34.69 
26  40 

Four  hours  after  adding  salt  

54.60 

3408 

37  59 

Five  hours  after  adding  salt  

54.00 

30.96 

42  65 

Six  hours  after  adding  salt  .         .    . 

57.06 

27.12 

52.48 

Seven  hours  after  adding  salt.     .    . 

50.52 

23.10 

54.28 

Eight  hours  after  adding  salt  

55.20 

20.04 

63.70 

(a)  The  charge,  subjected  for  4  hours  to  oxidizing  roasting  before  7  per 
cent,  salt  was  added,  yielded  64.27  per  cent,  of  the  silver  soluble  in  sodium 
hyposulphite.  The  maximum  decomposition  was  reached  in  3  hours  after 
the  salt  was  added,  at  which  time  58.92  per  cent,  of  the  soluble  silver  was 
rendered  insoluble.  The  decomposition  during  the  first  hour  was  14.81  per 
cent.;  the  second  hour,  31.21  per  cent.;  the  third  hour,  12.90.  During  the 
next  5  hours  we  find  the  amount  of  soluble  silver  gradually  increasing  until 
at  the  end  of  the  fifth  hour  nearly  the  same  amount  of  silver  was  rendered 
soluble  as  the  ore  contained  before  the  salt  was  added.  This  would  indicate 
that  by  a  continuation  good  results  may  finally  be  obtained,  but  the  heat  of 
the  charge  is  exhausted  before  that  time,  and  a  second  fire  causes  decomposi- 
tion again. 


CHLORIDIZING  OF  CALCAREOUS   ORES 


139 


CHARGE  NO.  125 

(4  per  cent,  salt  added  during  roasting.) 


VALUE  OF 

VALUE  OF 

ROASTED  ORE 

DESCRIPTION  (a) 

CONTAINING 

LEACH  TAIL- 

SOLUBLE 

SALT 

INGS 

SILVER 

PER  TON 

PER  TON 

Oz.  Silver 

Oz.  Silver 

Per  Cent. 

Raw  ore                   Oz  per  ton  59  70 

After  3  hours  roasting  and  just  be- 

fore adding  salt  .  .Oz.  per  ton,  58.32 

26.58 

54.42 

Just  after  adding  salt  and  stirred    .  . 

52.44 

20.84 

61.79 

One  hour  after  adding  salt  

56.16 

23.94 

57.38 

Two  hours  after  adding  salt  

56.10 

29.40 

47.60 

Three  hours  after  adding  salt  

54.90 

42.24 

23.07 

Four  hours  after  adding  salt  

58.56 

46.80 

20.09 

Five  hours  after  adding  salt  

56.50 

39.66 

29.25 

(a)  After  3  hours  oxidizing  roasting  54.42  per  cent,  of  the  silver  in  the  ore 
was  soluble  in  sodium  hyposulphite,  then  4  per  cent,  salt  was  added  and 
mixed  with  the  ore  by  hoes.  A  sample  then  taken  through  the  entire 
length  of  the  furnace  showed  61.79  per  cent,  soluble  silver,  or  a  gain  of  7.39 
per  cent,  in  these  few  minutes.  Decomposition  soon  set  in,  however,  and 
the  maximum  was  reached  in  four  hours  after  the  salt  was  added,  at  which 
time  67.48  per  cent,  of  the  soluble  silver  was  rendered  insoluble.  The 
decomposition  during  the  first  hour  was  7.13  per  cent;  the  second  hour,  15.83 
per  cent.;  the  third  hour,  39.70  per  cent.;  the  fourth  hour,  4.82  per  cent. 

Summing  up  the  results  of  these  experiments  it  was  clear 
that  (1)  at  a  very  early  stage  of  oxidizing  roasting  quite  a  high 
percentage  of  the  silver  was  converted  into  a  combination  (un- 
doubtedly silver  arsenate)  which  was  soluble  in  sodium  hypo- 
sulphite, and  which  seemed  to  resist  well  the  decomposing  action 
of  the  lime.  (2)  A  part  of  this  soluble  silver  was  decomposed  by 
the  action  of  the  salt,  the  decomposition  commencing  almost 
immediately  after  the  addition  of  salt  and  continuing  until  a 
maximum  was  reached,  when  a  reaction  took  place  by  which 
soluble  silver  was  again  formed  —  most  likely  silver  chloride. 
(3)  A  larger  percentage  of  salt  produced  the  decomposition 
quicker,  but  not  so  thoroughly  as  a  smaller  percentage,  viz. : 

Charge  No.  116,  with  10  per  cent,  salt,  reached  the  maximum 
in  2  hours,  with  51.92  per  cent,  of  soluble  silver  decomposed. 

Charge  No.  123,  with  7  per  cent,  salt,  reached  the  maximum  in 
3  hours,  with  64.27  per  cent,  of  soluble  silver  decomposed. 

Charge  No.  125,  with  4  per  cent,  salt,  reached  the  maximum 
in  4  hours,  with  67.48  per  cent,  of  soluble  silver  decomposed. 


140  HYDROMETALLURGY  OF  SILVER 

The  singular  fact  was  developed  that  when  salt  was  added  in 
the  reverberatory  furnace  during  the  oxidizing  period  no  such 
decomposition  took  place.  Lixiviating  tests  made  hourly  and 
half-hourly  after  the  salt  was  added  showed  a  gradually  increas- 
ing chlorination.  This  strange  behavior  of  the  ore  in  the  Bruck- 
ner seems  to  be  a  reaction  of  silver  arsenate  and  sodium  chloride, 
but  why  such  a  reaction  does  not  take  place  in  the  reverberatory 
I  cannot  explain.  No  lumps  or  balls  were  formed  in  this  mode 
of  roasting,  but  the  peculiar  reaction  prevented  a  higher  chlorina- 
tion and  rendered  this  method  of  roasting  impracticable. 

ADDING    THE    SALT   IN    THE    BATTERY.  —  SELF-ROASTING 

When  the  salt  is  added  in  the  battery  it  becomes  thoroughly 
mixed  with  the  ore,  and  both  enter  the  furnace  together.  Though 
there  is  also  a  considerable  part  of  the  silver  converted  into  a 
salt  soluble  in  sodium  hyposulphite  at  a  very  early  stage  of  roast- 
ing, due  to  the  arsenic  in  the  ore,  yet  no  decomposition  of  the 
arsenate  of  silver  seems  to  take  place.  On  the  contrary,  a  gradual 
increase  of  soluble  silver  is  observed  up  to  a  very  advanced  stage 
of  roasting,  when  all  the  sulphureted  minerals  are  converted  into 
oxides.  If  the  roasting  is  continued  beyond  this  point,  or  if 
the  temperature  is  raised  before  or  after  this  point  is  reached, 
then  the  caustic  lime  acts  on  the  silver  chloride  and  decomposes  it. 

When  beginning  the  experiments  I  conducted  the  roasting 
in  the  Bruckner  the  same  as  with  certain  highly  sulphureted 
silicious  ores,  viz.:  (1)  by  raising  a  strong  heat  to  start  combus- 
tion; (2)  oxidizing  without  fire  until  the  temperature  commenced 
to  decrease;  (3)  chloridizing  and  finishing  at  a  higher  tempera- 
ture with  a  second  fire.  In  the  course  of  the  experiments,  how- 
ever, the  employment  of  the  second  fire  caused  the  decomposition 
of  a  considerable  part  of  the  soluble  silver  already  formed.  A 
charge  was  therefore  roasted  without  using  a  second  fire,  leaving 
the  ore  to  complete  the  roasting  in  the  heat  created  by  the  com- 
bustion of  the  sulphurets,  and  much  better  results  were  attained. 
For  the  sake  of  comparison  I  roasted  and  finished  a  number  of 
charges  without  the  additional  fire  and  then  a  number  with  the 
second  fire,  using  ore  of  the  same  grade  and  maintaining  simi- 
larity in  all  other  conditions,  and  found  that  with  self-roasting 
the  average  chlorination  was  5  per  cent,  higher,  while  the  consump- 


CHLORIDIZING  OF  CALCAREOUS  ORES  141 

tion  of  wood  was  reduced  more  than  one-half.  With  a  second 
fire  the  average  value  of  the  roasted  ore  was  0.46  oz.  silver  less 
than  the  value  of  the  corresponding  raw  ore,  while  in  self-roasting 
the  average  value  of  the  roasted  ore  was  1.91  oz.  silver  higher 
than  that  of  the  raw  ore,  indicating  smaller  loss  by  volatilization. 
I  call  this  mode  of  roasting  "self-roasting,"  because  the  ore, 
once  ignited,  requires  no  further  attention,  and  is  left  entirely 
to  itself  until  the  heat  has  nearly  died  out.  Self-roasting  sim- 
plifies the  manipulation.  It  is  only  necessary  to  maintain  fire 
for  about  two  hours,  when  the  charge  can  be  left  to  itself  until 
the  time  of  discharging,  which  is  about  an  hour  before  the  red 
heat  dies  out.  Thus  one  man  can  attend  to  quite  a  number  of 
furnaces.  Care  has  to  be  taken,  however,  to  make  large  charges, 
as  small  charges  do  not  maintain  the  heat  long  enough,  and  they 
"freeze"  before  the  roasting  is  completed. 

Having  obtained  by  this  mode  of  roasting  an  average  of  82.9 
per  cent,  chlorination,  with  occasional  results  of  85  and  86  per 
cent,  (see  Table)  with  calcareous  ore  which  offered  so  many  dif- 
ficulties to  chloridizing  roasting,  it  seems  that  it  may  be  possible 
to  chloridize  properly  nearly  all  kinds  of  highly  sulphureted  ores 
in  this  way;  but,  as  large  charges  work  better  than  small  ones, 
it  is  advisable  to  have  the  revolving  cylinders  large  enough  to 
hold  5  or  6  tons  of  ore. 

By  roasting  in  this  manner  more  subchlorides  are  formed, 
and  less  volatile  chlorides  expelled;  this,  however,  does  not  inter- 
fere much  with  the  subsequent  lixiviation.  The  advantages 
gained  will  more  than  overbalance  the  slight  extra  expense 
caused  by  the  increased  consumption  of  sulphur,  and  the  refining 
of  a  somewhat  baser  precipitate,  especially  if  the  diminished  loss 
of  silver  by  volatilization,  which  loss  is  principally  caused  by  the 
expulsion  of  the  volatile  base-metal  chlorides,  is  taken  into  con- 
sideration. If  the  ore  contains  copper,  an  increased  formation 
of  cuprous  chloride  will  even  be  beneficial  for  the  subsequent 
extraction  by  lixiviation. 

The  consumption  of  wood  in  self-roasting  was  found  to  be 
only  one  cord  for  each  10.6  tons  of  ore,  while  in  roasting  with  a 
second  fire  only  4  to  4.5  tons  could  be  roasted  with  a  cord  of 
wood. 


142  HYDROMETALLURGY  OF  SILVER 

BALLING  OF  THE  ORE 

Another  noticeable  difference  in  the  behavior  of  the  ore,  due 
to  the  addition  of  the  salt  in  the  battery  or  in  the  furnace,  was 
the  formation  of  balls  or  lumps  in  the  first  case,  while  in  the  latter 
instance  the  ore  remained  loose  without  forming  balls,  and  when 
discharged  ran  on  the  cooling  floor  like  water.  When  the  salt 
was  added  in  the  battery  it  assumed  a  more  solid  form,  did  not 
spread  over  the  cooling  floor,  and  did  not  dust,  but  contained  a 
great  many  balls.  These  balls  originated  during  the  early  part 
of  the  chloridizing  period  and  at  first  were  of  the  size  of  a  pin's 
head.  Gradually  they  assumed  larger  dimensions,  and  when 
the  charge  was  finished  the  majority  of  them  were  from  the  size 
of  a  pea  to  that  of  a  walnut.  They  were  smooth,  hard,  and  heavy, 
consisting  of  concentric  shells,  and  were  formed  even  if  a  second 
fire  was  not  used.  By  reducing  the  speed  of  the  furnace  from 
2£  revolutions  per  minute  to  one  revolution  in  If  minutes,  and  by 
adopting  self-roasting,  which  was  equivalent  to  roasting  at  the 
lowest  possible  heat,  the  number  and  size  of  these  hard  balls 
were  greatly  reduced,  but  were  not  altogether  prevented. 

Roasting  during  the  chloridizing  period  was  tried  with  an 
intermittent  motion.  The  furnace  was  allowed  to  make  one 
revolution  and  was  then  stopped  for  fifteen  minutes,  when  another 
revolution  was  made,  and  so  on  until  the  charge  was  finished. 
This  reduced  but  did  not  prevent  the  formation  of  balls.  Then 
the  furnace  was  stopped  entirely  during  the  chloridizing  period, 
intending  to  allow  the  completion  of  the  roasting  process  with- 
out any  further  movement,  though  it  extended  the  time  required 
for  roasting.  This  method  would  very  likely  have  had  the  de- 
sired effect  had  not  another  difficulty  made  it  impracticable. 
The  surface,  of  the  ore  commenced  to  harden,  forming  a  crust, 
which  increased  in  thickness,  was  spongy  and  porous,  and  would 
not  have  interfered  with  the  lixiviation,  but  threatened  that  by 
the  time  the  charge  was  finished  the  whole  mass  might  have 
hardened,  which  would  have  caused  great  difficulty  in  discharg- 
ing, and  would  also  have  endangered  the  furnace  through  the 
whole  mass  clinging  to  one  side  and  then  dropping  suddenly  as 
the  furnace  revolved. 

Though  annoying,  this  formation  of  hard  balls  would  not  be 
a  serious  obstacle,  as  by  repeated  tests  it  was  found  that  they 


CHLORIDIZING  OF  CALCAREOUS   ORES  143 

were  well  roasted;  in  fact,  always  a  slight  percentage  better 
chloridized  than  the  fine  stuff.  By  separating  the  coarse  from 
the  fine,  and  crushing  the  former  in  a  ball-mill  or  through  rolls, 
the  ore  would  be  well  prepared  for  lixiviation.  This  extra  hand- 
ling does  not  signify  much  when  the  great  difference  in  expense 
between  roasting  in  a  Bruckner  and  roasting  in  a  reverberatory 
furnace  is  taken  into  consideration,  to  say  nothing  of  other 
advantages.  Other  lumps  were  formed  toward  the  end  of  the 
operation  which  were  larger,  but  soft  and  porous,  had  a  rough 
surface,  fell  apart  when  brought  in  contact  with  water,  and  did 
not  interfere  with  lixiviation. 

The  sulphureted  minerals  of  this  ore  had  no  tendency  to  cake 
and  form  lumps.  When  the  concentrates,  free  from  gangue, 
were  roasted,  the  pulp  remained  perfectly  loose  and  sandy,  even 
if  only  occasionally  stirred  during  roasting;  the  formation  of  these 
small  hard  balls  must  therefore  have  been  caused  by  the  gangue. 
They  were  soft  while  hot  and  were  readily  crushed,  but  when 
cold  became  hard  and  brittle,  and  when  broken  showed  concentric 
layers.  They  were  not  caused  by  excessive  heat,  because  they 
formed  even  when  the  heat  was  kept  as  low  as  the  combustion 
of  the  sulphureted  minerals  permitted;  neither  was  their  appear- 
ance that  of  overheated  ore.  They  were  probably  caused  by  the 
formation  of  a  double  salt  of  calcium  sulphate  and  sodium  sulphate 
(glauberite),  Na2SO4  -f-  CaSO4,  a  salt  which  fuses  easily.  This 
would  explain  the  singular  fact  that  no  lumps  were  formed  if  the 
salt  was  added  in  the  furnace  after  the  charge  had  been  oxidizing 
for  some  time;  in  which  case  the  salt  could  not  be  thoroughly 
mixed  with  the  ore,  and  therefore  did  not  come  in  such  close 
contact  with  the  sulphureted  matters  as  the  lime  did,  on  which 
sulphuric  acid  acts  so  much  more  energetically  than  on  sodium 
chloride  that  very  little,  if  any,  of  the  sodium  chloride  was  con- 
verted into  sulphate.  For  the  same  reason  free  chlorine  could 
be  detected  only  in  exceptional  cases  if  the  salt  was  added  in  the 
furnace.  Neither  was  it  possible  to  produce  a  chlorination  of 
the  silver.  The  40  to  60  per  cent,  soluble  silver  was  arsenate  of 
silver. 

In  my  muffle  experiments  with  different  mixtures  of  concen- 
trates and  gangue  (limestone)  the  tendency  to  form  lumps  com- 
menced with  the  proportion  of  50  concentrates  to  50  gangue, 
and  this  tendency  increased  with  the  percentage  of  gangue;  it  did 


144  HYDROMETALLURGY  OF  SILVER 

not  exist  if  the  mixture  contained  more  sulphurets  than  gangue, 
and  if  the  ore  had  been  assorted  at  the  mine  to  meet  this  require- 
ment the  formation  of  balls  would  doubtless  have  been  entirely 
avoided,  the  chlorination  would  have  been  much  better,  and  the 
Bruckner  furnaces  could  have  been  used  instead  of  the  more 
expensive  reverberatories.  It  may  be  mentioned  that  the  ore 
from  the  mine  had  to  pass  successively  through  three  large  ore- 
bins  before  reaching  the  mill.  These  were  always  kept  full,  in 
order  to  have  ore  in  reserve,  and  it  took  10  to  14  days  from  the 
time  the  ore  was  dumped  into  the  bin  at  the  mine  before  it  reached 
the  battery;  quick  changes  for  experimental  purposes  were 
therefore  impossible. 

ROASTING  IN  THE  REVERBERATORY  FURNACES 

There  were  in  all  eight  reverberatory  furnaces  in  operation, 
of  the  following  dimensions: 

Four  furnaces,  50  ft.  long  by  10  ft.  wide,  containing  5  hearths  each. 
Two  furnaces,  40  ft.  long  by  10  ft.  wide,  containing  4  hearths  each. 
Two  furnaces,  30  ft.  long  by  10  ft.  wide,  containing  3  hearths  each. 

The  arch  at  the  highest  place  near  the  fire  was  27  in.  above 
the  hearth,  further  away  only  20  in.,  and  at  the  last  hearth  18  in. 
The  sides  were  10  in.  high.  The  furnaces  were  built  in  pairs, 
placed  back  to  back,  and  each  hearth  had  but  one  small  working 
door  8  x  12  in. 

While  tolerably  good  chlorination  may  be  obtained  with 
silicious  silver  ore  in  furnaces  of  the  above-described  construction, 
they  surely  were  not  of  suitable  design  for  the  roasting  of  calca- 
reous silver  ores,  as  in  the  first  place  insufficient  provision  was 
made  for  a  free  and  well-located  air  inlet;  the  small  working  door 
furnished  air  only  for  the  ore  near  it,  which  fumed,  while  further 
in  and  toward  the  back  of  the  furnace  it  presented  the  appearance 
of  an  inactive  glowing  mass  of  high  temperature,  and  did  not 
emit  any  visible  fumes.  In  the  second  place,  the  arch,  especially 
the  one  over  the  hearth  nearest  to  the  fire,  was  not  high  enough 
to  keep  the  temperature  low  enough  for  calcareous  ores  during 
the  chloridizing  period.  There  was  no  remedy  for  these  defects 
except  reconstruction.  By  keeping  the  fire-door  continually  wide 
open  much  better  results  were  obtained,  but  they  were  not 
satisfactory  until  the  furnaces  were  reconstructed. 


CHLORIDIZING  OF  CALCAREOUS  ORES  145 

At  Yedras  I  was  astonished  to  find  that  the  ore  in  the  rever- 
beratory  furnaces  on  the  finishing  hearth  was  subjected  to  almost 
a  white  heat,  while  care  was  taken  to  avoid  cooling  the  furnace 
by  excluding  the  air  as  much  as  possible.  The  fire-door  was 
kept  closed,  and  the  very  small  working  doors,  of  which  there  was 
one  on  each  hearth,  were  closed  as  soon  as  stirring  was  completed. 
One  of  the  charges  roasted  in  this  way  consisted  of  one  ton  of 
ore;  four  men  were  attending  the  furnace,  which  was  50  ft.  long, 
and  5  per  cent,  salt  was  added  on  the  third  hearth  after  the  ore 
had  been  nearly  five  hours  in  the  furnace.  The  following  is  a 
detailed  record  of  same: 

1.  Raw  ore  without   salt.     Sample  of  ore  taken  from  the 
hopper  before  charging  gave  67.08  oz.  silver  per  ton. 

2.  Same  charge.     Sample  taken  from  the  first  hearth  after 
the  ore  had  been  two  hours  in  the  furnace,  and  just  before  the 
charge  was  moved  to  the  second  hearth,  the  hearth  being  dark, 
gave  off  a  pretty  strong  smell  of  sulphurous  acid  and  contained 
65.40  oz.  silver  per  ton  of  ore. 

3.  Sample  taken  from  the  second  hearth,  after  having  been 
there  for  two  hours  and  twenty  minutes,  and  just  before  the 
charge  was  to  be  moved  to  the  third  hearth.     Charge  dark-red, 
strong  fume  (arsenic  period),  with  strong  smell  of  sulphurous  acid, 
gave  66.06  oz.  silver  per  ton  of  ore. 

4.  Sample  taken  from  the  third  hearth  just  before  the  salt 
was  added.     Charge  red-hot,  strong  smell  of  sulphurous  acid, 
gave: 

Leach  tailings  31  44  oz   |  53. 08  per  cent,  silver  soluble  in  sodium  hyposulphite. 

5.  Sample  taken  from  third  hearth  after  having  roasted  just 
one  hour  with  salt.     Strong  fumes  and  smell  of  sulphurous  acid; 
color,  after  cooling,  light-brown;  gave: 

Leach  tailings  29.04  oz!  |  54'4  Percent"  silver  soluble  in  sodium  hyposulphite. 

6.  Sample  taken  from  fourth  hearth  after  the  charge  had 
been  there  one  hour,  and  had  roasted  two  hours  with  salt.     The 
ore  woolly,  and  sample  exposed  to  air  fumed  strongly;  color,  after 
cooling,  red-brown;  temperature  red;  gave: 

Leach  tailings  28'.44  oz'.  |  53  Per  cent  silver  soluble  in  sodium  hyposulphite. 


146  HYDROMETALLURGY   OF  SILVER 

7.  Sample  taken  while  charge  was  being  moved  to  fifth  hearth 
and  after  roasting  three  hours  and  twenty  minutes  with  salt. 
Temperature  light  red;  sample  fumed  strongly  when  exposed  to 
air;  inside  of  the  furnace  clear,  fumed  only  near  the  working  door; 
slight  smell  of  chlorine  and  sulphurous  acid;  ore  commenced  to 
assume  a  sandy  consistency,  and  had  the  appearance  of  an  over- 
heated ore;  color,  after  cooling,  red-brown;  gave: 

Lelch  tailings  27.12  oz'.  }  61'8  Per  cent"  silver  soluble  in  sodium  hyposulphite. 

8.  Sample  taken  from  fifth  hearth  after  roasting  five  hours 
with  salt.     No  smell  of  sulphurous  acid,  very  little  of  chlorine; 
temperature  very  light  red,  almost  white;  ore  sandy;  color,  after 
cooling,  greenish  brown,  gave: 

Lelch  tailings  23'.94  oz!  |  65'  2  per  cent>  silver  soluble  in  sodium  hyposulphite. 

9.  Sample  taken  when  the  ore  was  discharged,  after  having 
been  eleven  hours  in  the  furnace  and  six  hours  roasted  with  salt. 
Very  little  smell  of  chlorine,  none  of  sulphurous  acid;  ore  very 
sandy,  mixed  with  lumps;  color,  after  cooling,  greenish  brown; 
gave: 

Leach  tailings  2L06  oz!  }  67'2  Per  cent  silver  soluble  in  sodium  hyP°sulphite. 

The  ore  sustained  during  roasting  a  loss  in  weight  by  vola- 
tilization of  13.8  per  cent.,  and  the  loss  of  silver  by  volatilization 
amounted  to  16  per  cent. 

By  comparing  the  different  samples  it  was  found  that  at  the 
time  the  salt  was  added  there  was  53.8  per  cent,  of  the  silver 
soluble  in  sodium  hyposulphite.  This  soluble  silver  was  an 
arsenate  and  was  formed  in  the  same  manner  as  that  obtained, 
in  the  Bruckner  furnace  during  the  arsenic  period.  During  the 
next  two  hours  of  roasting  with  salt  there  was  no  change  in 
the  soluble  silver.  This  is  unlike  the  behavior  of  the  ore  in  the 
Bruckner,  in  which  during  this  period  a  marked  decomposition 
of  the  soluble  silver  took  place.  Although  in  the  reverberatory 
the  salt  was  also  added  after  the  ores  had  been  oxidizing  in  the 
furnace  for  nearly  five  hours,  and  the  arsenic  period  was  over, 
no  soluble  silver  was  decomposed  by  the  salt.  During  the  next 
two  hours  it  remained  the  same,  and  commenced  to  increase  only 
after  the  third  hour  with  salt,  and  after  the  ore  had  reached  the 


CHLORIDIZING  OF  CALCAREOUS  ORES  147 

region  of  light-red  heat  on  the  fourth  hearth.  The  increase, 
however,  amounted  to  only  8  per  cent,  and  on  the  finishing 
hearth,  where  the  ore  was  exposed  to  an  almost  white  heat  for 
two  hours  and  forty  minutes,  a  further  increase  of  5.4  per  cent, 
occurred  —  in  all  13.4  per  cent.  —  while  the  loss  of  silver  by 
volatilization,  which  took  place  principally  during  this  period, 
amounted  to  16  per  cent.  Of  the  67.2  per  cent,  of  silver  rendered 
soluble  in  sodium  hyposulphite,  53.8  per  cent,  was  due  to  the 
action  of  the  arsenic,  and  only  13.4  per  cent,  to  the  action  of  the 
salt.  To  gain  13.4  per  cent,  chlorination,  16  per  cent,  of  silver 
was  sacrificed,  a  very  thoughtless  operation;  and  this  was  by  no 
means  one  of  the  worst  results.  With  most  of  the  charges  25  to 
30  per  cent.,  and  even  more,  of  the  silver  was  lost,  to  gain  a  small 
percentage  in  chlorination,  as  is  shown  further  on. 

The  low  chlorination  was  due  to  the  insufficient  supply  of  air, 
which  prevented  the  formation  of  sulphuric  acid  to  satisfy  the 
lime  and  to  act  on  the  salt,  while  the  great  loss  of  silver  was 
caused  by  excessive  heat  and  insufficient  air.  The  different 
behavior  of  the  ore  in  the  two  furnaces  as  regards  the  decompo- 
sition of  the  soluble  silver  after  the  salt  was  added  is  remarkable, 
and  is  thus  far  unexplainable. 

While  experimenting  with  the  Bruckner  furnaces,  the  roasting 
in  the  reverberatory  furnaces  was  conducted  in  the  manner  just 
described,  and  the  experiments  with  these  commenced  after  those 
with  the  Bruckners  were  completed.  One  preliminary  experiment, 
however,  was  made,  to  demonstrate  to  the  metallurgist  in  charge 
the  effect  of  the  air;  for  he  insisted  that  the  Yedras  ore  could 
only  be  chloridized  at  a  very  light-red  heat,  and  he  excluded  the 
air  as  far  as  possible,  so  as  not  to  cool  the  furnace. 

As  mentioned  above,  the  furnaces  were  not  of  proper  construc- 
tion, and  the  only  means  of  getting  more  air  into  them  was  by 
leaving  the  fire-door  wide  open.  The  beneficial  effect  of  the  air 
was  quite  striking,  though  no  change  was  made  in  the  temperature, 
the  roasting  proceeding  at  the  same  high  heat  as  before.  The 
interesting  record  is  given  in  the  following  tables: 


148 


HYDROMETALLURGY  OF  SILVER 


ROASTING    WITH   5   PER  CENT.  SALT  AT  A  VERY  HIGH  HEAT 

WITH  CLOSED  FIRE-DOOR 

Reverberatory  Furnace,  No.  3 


DATE 

VALUE  OF 
RAW  ORE  PER 
TON    (a) 

VALUE  OF 
ROASTED  ORE 
PER  TON 

SOLUBLE 
SILVER 

Loss  OF  SILVER 
BY  VOLATILIZA- 
TION 

January  13  
14 

Oz.  Silver 
63.03 
62.04 

Oz.  Silver 
56.52 
50.04 

Per  Cent. 
67.7 
696 

Per  Cent. 

17.8 
25  5 

15  
16  
17  
18  
19  
20  
21 

63.78 
63.78 
64.08 
61.20 
60.60 
61.56 
60.00 

49.44 
52.80 
60.00 
49.08 
45.00 
49.50 
48.48 

57.1 
49.8 
43.0 
43.2 
63.2 
68.8 
68  1 

27.2 
25.4 
13.9 
26.3 
31.7 
26.1 
25  7 

22 

58.20 

45.48 

76  8 

278 

Average  

61.82 

50.6*3 

60.7 

24.7 

(a)  Difference  of  value  between  raw  and  roasted  ore,  11.9  oz.  per  ton. 

ROASTING  WITH   5  PER  CENT.  SALT   AT  A  VERY  HIGH  HEAT 
WITH  OPEN  FIRE-DOOR 
Reverberatory  Furnace,  No.  3 


VALUE  OF 

VALUE  OF 

Loss  OF  SILVER 

DATE 

RAW  ORE  PER 

ROASTED  ORE 

BY  VOLATILIZA- 

TON. (a) 

PER  TON 

SILVER 

TION 

Oz.  Silver 

Oz.  Silver 

Per  Cent. 

Per  Cent. 

January  23  

57.66 

49.56 

85.3 

20.6 

24  

55.50 

47.04 

79.7 

21.6 

25  

54.36 

50.40 

79.6 

14.3 

26  

63.30 

55.20 

83.9 

18.0 

27  

63.84 

59.22 

77.7 

13.9 

28  

64.86 

61.08 

73.5 

12.6 

29  

65.52 

57.00 

65.5 

19.2 

30  

66.78 

59.40 

77.2 

17.4 

31  

66.00 

58.80 

74.5 

17.3 

February  1  

62.10 

55.68 

79.6 

16.7 

2  

57.30 

53.40 

75.7 

13.7 

3  

54.54 

48.96 

76.2 

16.6 

4  

55.20 

52.80 

77.5 

11.2 

5  

57.60 

51.96 

75.8 

12.0 

6  

60.66 

55.14 

80.5 

15.6 

7  

61.68 

53.70 

72.1 

19.2 

8  

64.50 

65.64 

68.9 

5.5 

9  

55.80 

55.56 

75.5 

7.6 

10 

57.96 

48.48 

80.7 

22.3 

11  

62.46 

56.52 

76.8 

16.0 

12  

67.02 

59.94 

69.5 

10.7 

Average  

60.70 

55.02 

76.4 

15.3 

(a)  Difference  of  value  between  raw  and  roasted  ore,  5.68  oz.  per  ton. 


CHLORIDIZING  OF  CALCAREOUS  ORES  149 

By  comparing  the  two  tables  we  find  that  from  the  very  first 
day  the  fire-door  was  kept  open  a  marked  improvement  in  the 
results  took  place.  The  averages  show  this  plainly: 

Per  Cent. 

Average  loss  of  silver — fire-door  closed 24.7 

Average  loss  of  silver — fire-door  open 15.3 

Difference  in  favor  of  open  fire-door 9.4 

Per  Cent. 

Average  of  soluble  silver  —  fire-door  open 76.4 

Average  of  soluble  silver — fire-door  closed 60.7 

Difference  in  favor  of  open  fire-door 15.7 

Thus  a  reduction  of  9.4  per  cent,  in  the  loss  of  silver  and  an 
increase  of  15.7  per  cent,  in  soluble  silver,  or  a  total  saving  of 
25.1  per  cent,  of  the  silver  in  the  ore,  was  effected  by  simply 
admitting  more  air  into  the  furnace. 

Preparatory  to  my  experiments  with  the  reverberatory  fur- 
naces I  reconstructed  some  of  them,  to  adapt  them  to  the  require- 
ments of  the  ore.  The  arch  was  raised  to  3  ft.  with  the  exception 
of  the  hearth  furthest  from  the  fire  on  which  the  ore  was  charged 
and  which  was  much  lower.  The  working  doors  and  the  flue 
were  enlarged,  and  air-channels  constructed  to  permit  air  to  enter 
along  the  fire-bridge  and  along  the  back  wall  of  the  furnace. 

Roasting  was  then  commenced  at  a  much  lower  temperature, 
and  with  a  more  liberal  supply  of  air,  much  better  results  being 
attained.  The  average  chlorination  of  one  month  rose  to  81.7 
per  cent,  while  the  loss  of  silver  was  reduced  to  1.7  per  cent. 
In  the  new  furnaces  the  ore  fumed  over  the  whole  hearth  instead 
of  only  near  the  working  doors  as  before,  and  the  temperature 
was  much  more  uniform  in  all  parts  of  each  hearth.  To  roast  at 
a  low  heat  requires  much  greater  attention  and  skill  than  at  a 
high  heat,  and  it  took  some  time  before  the  men  became  accus- 
tomed to  it. 

After  numerous  experiments  to  determine  the  proper  tempera- 
ture and  draft,  the  following  rules  were  adopted  and  will  apply  to 
all  calcareous  sulphureted  silver  ores: 

1.  The  fumes  evolved  must  be  kept  in  motion  in  all  parts  of 
the  furnace.  If  they  stagnate  around  the  ore,  or  if  the  furnace 
assumes  nearly  a  uniform  heat  throughout  its  entire  length,  it  is 
always  a  sign  of  insufficient  draft,  and  if  the  draft  is  not  increased 
the  result  will  invariably  be  a  high  loss  of  silver  and  a  low  chlori- 
nation. 


150  HYDROMETALLURGY  OF  SILVER 

2.  In  no  part  of  the  furnace  should  the  ore  attain  a  light-red 
heat.  The  fire  should  be  regulated  entirely  according  to  the 
temperature  required  at  the  finishing  hearth  nearest  the  fire, 
where  the  temperature  should  be  kept  so  that  the  ore  has  rather 
a  dark  surface  if  not  stirred,  but  a  dark-red  heat  when  the  hoe 
enters  it.  The  ore,  however,  should  fume,  and  the  temperature 
never  go  below  the  point  required  to  evolve  fumes.  The  presence 
of  fumes  always  indicates  that  the  temperature  is  not  too  low, 
while  the  dark  surface  of  the  ore  shows  that  the  temperature  is 
not  too  high. 

An  arch  still  higher  than  3  ft.  facilitates  the  maintenance  of 
the  conditions  prescribed  in  the  second  rule,  and  at  the  same 
time  permits  a  stronger  fire  for  the  more  remote  hearths  without 
injuring  the  charge  on  the  finishing  hearth.  I  rebuilt  furnaces 
Nos.  3  and  4  and  made  the  arch  of  the  finishing  hearth  5  ft.  high, 
the  sides  3  ft.  4  in.,  and  the  fire-bridge  2  ft.  high;  the  next  two 
hearths  were  raised  so  that  the  arch  was  only  3  ft.  high,  the  sides 
16  in.;  the  next  arch  27  in.  high,  sides  16  in.;  and  the  charging- 
hearth  arch  24  in.  high  and  sides  14  in.  (Figs.  4,  5  and  6).  These 
gave  the  best  and  most  uniform  results  of  any  of  the  furnaces; 
the  proper  temperature  on  the  finishing  hearth  being  maintained 
much  easier,  because  the  flame  following  the  roof  was  so  far 
above  the  ore  that  it  would  have  taken  an  excessive  fire  to  over- 
heat the  charge.  The  chlorination  obtained  was  very  satisfactory, 
and  the  loss  of  silver  by  volatilization  was  reduced  to  a  minimum. 
The  average  results  of  six  weeks'  working  were:  chlorination, 
83.8  per  cent;  loss  of  silver,  0.8  per  cent. 

The  consumption  of  wood  in  these  furnaces  was  larger  than 
in  the  others:  in  the  reverberatories  with  3-ft.  arch  it  amounted 
to  0.098  cord  per  ton  of  ore  roasted;  and  in  the  furnaces  with  a 
5-ft.  high  arch  0.32  of  a  cord  more  per  ton  of  ore  was  used;  but 
the  better  results,  and  the  ease  with  which  the  temperature  could 
be  controlled,  overbalanced  the  extra  consumption  of  wood. 

When  the  roasting  was  properly  conducted  the  percentage  of 
chlorination  invariably  depended,  as  in  the  Bruckner  furnaces, 
upon  the  proportion  of  sulphureted  minerals  and  calcareous 
gangue  of  which  the  ore  was  composed.  The  more  sulphureted 
matters  the  ore  contained  the  better  was  the  result.  But  little 
difference  was  noticed  in  the  formation  of  lumps  whether  the 
salt  was  added  in  the  battery  or  in  the  furnace;  in  both  cases 


CHLORIDIZING  OF  CALCAREOUS  ORES  151 

lumps  were  formed,  but  as  a  rule  they  were  more  porous  and 
softer  than  those  of  the  Bruckner,  and  could  easily  be  leached. 


CONCLUSIONS 

My  experiments  in  roasting  the  calcareous  and  arsenical  ore 
of  Yedras  have  shown: 

1.  That  the  main -difficulty  in  chloridizing  the  ore  to  a  high 
percentage  was  caused  by  the  excess  of  lime  in  the  ore.     The  ore 
at  Yedras  was  assorted,  as  in  all  silver  mines,  with  the  view  to 
obtaining  a  certain  grade  in  silver,  and  no  attention  was  paid  to 
the  relative  proportion  of  lime  and  sulphureted  minerals,  which 
was  of  so  great  importance.     In  order  to  obtain  regular  chlori- 
nations  of  over  90  per  cent,  it  was  necessary  that  the  ore  should 
contain  not  less  than  50  per  cent,  of  sulphurets;  but  to  accomplish 
this  it  was  necessary  to  break  the  ore  smaller,  so  that  more  of 
the  barren  limestone  and  calcspar  could  be  thrown  out,  and 
enable  the  sorters  to  pick  out  a  poor  grade  of  second-class  ore. 
Pieces  containing  much  iron  pyrites,  even  if  they  did  not  contain 
silver,  should  be  thrown  with  the  first-class  ore. 

The  accumulation  of  a  second-class  ore  was  not  very  desirable 
at  Yedras,  because  it  could  only  be  treated  by  concentration,  and 
the  water  supply  was  very  limited,  except  during  the  rainy 
season.  However,  I  consider  it  more  rational  to  accumulate  a 
second-class  ore  dump  at  the  mine  for  future  treatment  than  a 
rich  tailings  pile  at  the  mill,  because  it  is  very  difficult  to  extract 
the  silver  from  tailings,  while  the  chances  are  that  after  further 
development  the  mine  may  produce  more  water;  but  even  if  it 
should  not,  the  accumulation  of  second-class  ore  would  not  be 
more  than  could  be  concentrated  during  the  rainy  season.  This 
question  is  a  very  important  one  and  due  attention  should  be 
paid  to  it.  The  outlay  would  be  comparatively  small  compared 
with  the  large  amount  of  silver  lost  in  the  tailings. 

By  improvements  in  the  mode  of  roasting  I  succeeded  in 
increasing  the  chlorination  from  65  to  81  per  cent,  and  above, 
and  in  reducing  the  loss  of  silver  from  25  to  1.7  per  cent,  and  less, 
thereby  nearly  doubling  the  production.  If  the  ore  had  been 
properly  sorted  the  increase  would  have  been  at  least  10  per  cent, 
more. 

2.  I  proved  that  such  ores  could  be  roasted  with  a  trifling 


152  HYDROMETALLURGY  OF  SILVER 

loss  of  silver  by  volatilization  if  the  roasting  was  conducted  as 
described  above,  and  that  the  enormous  loss  which  the  ore  for- 
merly sustained  was  caused  by  too  high  a  temperature  and  an 
insufficient  supply  of  air. 

3.  I  demonstrated  that  such  ores,  properly  assorted,  could  be 
successfully  and  very  cheaply  roasted  in  the  Bruckner  furnace 
by  the  self-roasting  process,  if  the  salt  is  added  to  the  ore  in  the 
battery,  provided  suitable  provisions  be  made  to  separate  and 
pulverize  the  hard  but  well-chloridized  balls  which  form  in  these 
furnaces. 


PART   II 
EXTRACTION  OF  THE  SILVER 


XIII 

LIXIYIATION  WITH  SODIUM  HYPOSULPHITE 

PERCY  and  Hauch  were  the  first  who  proposed  to  convert  the 
silver  in  ores  into  chloride  by  roasting  with  salt  and  to  extract 
the  silver  with  a  solution  of  sodium  hyposulphite,  but  Von  Patera 
was  the  first  to  apply  this  method  in  actual  practice  by  treating 
with  it  the  rich  and  complex  silver  ores  of  Joachimsthal,  Bohemia. 
These  ores  were  formerly  treated  by  amalgamation  and  also  by 
the  Augustin  method,  which  both  required  chloridizing  roasting; 
but,  on  account  of  the  great  loss  of  silver  which  these  ores  sus- 
tained during  chloridizing,  only  the  poorer  grades  were  worked 
by  these  methods.  Patera,  however,  changed  the  mode  of 
roasting,  inasmuch  as  he  applied  steam,  whereby  the  loss  of  silver 
by  volatilization  was  much  reduced,  and  made  it  possible  that 
even  the  richest  varieties  of  the  Joachimsthal  ore  could  be  suc- 
cessfully chloridized.  By  using  sodium  hyposulphite  as  solvent 
he  introduced  a  process  much  cleaner  and  superior  to  either  the 
amalgamation  or  Augustin  method.  It  was,  however,  executed 
on  a  rather  small  scale;  the  roasting  was  done  in  charges  of  500 
to  600  lb.,  and  the  lixiviation  was  performed  in  tubs  holding 
about  200  lb.  of  roasted  ore. 

This  process  was  introduced  at  Joachimsthal  in  1858.  Ten 
years  later,  in  the  fall  of  1868,  I  introduced  it  successfully  at 
La  Dura,  Sonora,  Mexico,  on  a  large  scale,  and  in  the  following 
years  at  Trinidad,  Las  Bronzas,  San  Marcial,  in  Sonora,  Mexico; 
then  at  Silver  King,  Arizona;  Monitor,  California;  Cusihuiri- 
achic,  Chihuahua,  Mexico,  and  at  other  places. 

In  course  of  time,  as  experience  with  different  ores  was  acquired, 
the  process  was  much  improved  chemically,  as  well  as  with  re- 
gard to  appliances. 

The  principle  on  which  this  process  is  based  is  the  property 
of  silver  chloride  of  being  insoluble  in  water,  while  it  dissolves 

155 


156  HYDROMETALLURGY  OF  SILVER 

readily  in  a  solution  of  sodium  hyposulphite,  even  if  the  solution 
is  very  dilute.  To  convert  the  silver,  which  in  nature  mostly 
occurs  as  sulphide,  into  chloride,  the  sulphide  ore  is  roasted  with 
salt  (sodium  chloride),  as  wre  have  seen  in  the  first  part  of  this 
treatise.  The  solution  dissolves  only  that  part  of  the  silver 
which  is  converted  into  chloride,  and  the  extraction  depends, 
therefore,  entirely  upon  the  quality  of  the  roasting.  The  chlori- 
dizing  roasting  is  undoubtedly  the  most  important  part  of  the 
process;  and  it  cannot  be  too  strongly  impressed  upon  the  minds 
of  operators  to  pay  their  principal  attention  to  this  operation, 
and  to  make  a  thorough  study  of  the  behavior  and  characteristics 
of  their  respective  ores. 

Before  going  into  the  details  of  the  different  operations  it 
will  be  more  instructive  to  give  a  short  description  of  the  process. 
In  the  first  part  of  this  treatise  we  have  seen  that,  if  a  complex 
ore  be  subjected  to  chloridizing  roasting,  not  only  the  silver 
sulphide  but  also  the  other  metal  sulphides  are  converted  into 
chlorides,  subchlorides,  sulphates  and  oxides.  A  number  of  these 
salts  are  soluble  in  water  and  have  to  be  removed  from  the  roasted 
ore  by  leaching  with  water  before  silver  extraction,  in  order  to 
prevent  them  from  entering  the  silver  solution  and  afterward 
the  silver  precipitate.  This  operation  is  termed  "base-metal 
leaching."  The  leaching  with  water,  however,  dissolves  also 
undecomposed  salt,  if  such  should  be  present,  and  the  sodium 
sulphate  which  was  formed  in  roasting. 

If  by  test  it  is  ascertained  that  the  soluble  chlorides  and  sul- 
phates have  been  removed  from  the  ore,  a  diluted  solution  of 
sodium  hyposulphite  is  applied  on  top  of  the  ore.  This  solution 
by  descending  through  the  ore  dissolves  the  silver  chloride. 
This  operation  is  termed  "silver  leaching."  The  outflowing 
solution  is  collected  in  special  tanks  (precipitation  tanks).  To 
this  silver  solution  a  solution  of  calcium  or  sodium  sulphide  is 
added,  by  which  the  silver  and  some  other  metals  present  are 
precipitated  as  sulphide.  By  agitating  the  solution  the  precipi- 
tate collects  in  dark-brown  flakes,  which  settle  quickly  to  the 
bottom  of  the  tank.  When  settled  the  clear  solution  is  decanted 
and  returned  to  the  process,  and  is  used  over  and  over  again  in- 
definitely. The  precipitate  is  filtered,  washed,  dried  and  charged 
into  a  small  roasting  furnace  to  burn  off  the  excess  of  sulphur 
which  it  contains.  This  is  done  at  a  very  low  heat,  and  only  as 


'   . 
LIXIVIATION   WITH  SODIUM  HYPOSULPHITE  157 

long  as  the  blue  sulphur  flame  lasts.  This  free  sulphur  in  the 
precipitate  is  derived  from  the  precipitant.  The  calcium  and 
sodium  sulphides  are  both  used  as  polysulphides.  In  precipitation 
only  one  atom  of  sulphur  combines  with  the  metal,  while  the 
balance  drops  down  as  such. 

The  calcined  silver  precipitate  is  finally  melted  on  a  lead- 
bath  in  a  cupeling  furnace  and  refined,  or  it  is  shipped  and 
sold  to  smelting  works. 

After  this  synopsis  of  the  process,  I  will  proceed  to  describe 
and  discuss  the  different  operations  and  appliances. 

BASE-METAL  LEACHING 

The  roasted  ore  is  charged  into  wooden  tanks,  of  which  each 
is  provided  with  a  filter  bottom.  The  size  of  these  tanks  depends 
on  the  intended  working  capacity  of  the  works,  the  filtering 
quality  and  the  chemical  nature  of  the  roasted  ore.  Deep  tanks 
should  never  be  used;  in  fact,  the  rule  should  be  observed  not  to 
make  them  deeper  than  5  ft.,  and  the  diameter  not  to  exceed 
16  to  20  ft.  Tanks  with  too  large  a  diameter  are  difficult  to  dis- 
charge. The  reason  why  tanks  should  not  be  made  deeper  than 
5  ft.  will  be  given  below. 

Construction  of  the  Leaching  Tank.  —  Fig.  31  represents  a 
vertical  'and  Fig.  32  a  horizontal  section  of  a  leaching  tank  16  ft. 
in  diameter  and  5  ft.  deep,  made  of  3-in.  clear  lumber.  To  use 
lighter  lumber  than  3  in.  is  not  an  economical  policy.  The 
lumber  should  be  red  wood  or  cedar,  preferably  the  former, 
because  it  is  more  indifferent  to  the  action  of  chemicals  than 
any  other  soft  wood.  The  bottom  can  be  made  of  12-in.  plank, 
but  the  staves  ought  not  to  be  made  over  6  in.  wide,  in  order  to 
facilitate  binding  by  the  hoops.  It  is  very  important  that  the 
tank  should  be  tight  in  order  not  to  lose  by  leakage  valuable 
silver  solution,  and  therefore  it  should  be  made  very  carefully. 
From  the  factory  the  parts  are  furnished  in  crated  bundles  and 
the  edges  suffer  considerably  in  transportation;  besides,  these 
edges  have  usually  only  a  rough  saw-finish,  and  in  putting  the 
tank  together  they  ought  to  be  planed  by  a  skilled  carpenter. 
For  this  purpose  several  extra  staves  are  always  sent  with  the 
shipment.  The  bottom  planks  should  be  kept  in  position  by 
wooden  pins.  Before  erecting  the  tank  the  timber  on  which  it 
is  going  to  set  should  be  prepared,  then  the  bottom  planks  placed 


158 


HYDROMETALLURGY  OF  SILVER 


on  top  of  it  and  pinned  together,  then  the  staves  with  their  groove 
slipped  into  the  edge  of  the  bottom  and  driven  tight.  The  groove 
in  the  staves  should  be  f  in.  deep.  The  tank  with  its  weight  has 
to  rest  on  a  circular  bed  of  timber  specially  prepared,  which  must 
be  thick  enough  to  leave  the  staves  clear  and  free  from  weight. 
Timbers  6x6  in.  will  answer;  but  they  ought  not  to  be  further 
apart  than  12  in.,  as  shown  in  Fig.  31.  The  best  hoops  are 
those  made  in  two  or  three  sections  of  J  or  1-in.  round  iron,  and 


! 


I 


/1'Qrove 
ll*  Dressed 


^1* ^ 


>"£ 


1 


Wro|ught 
IroiJ  Hoo 


1&  Pipe 


FIG.  31.  — LEACHING  TANK,  VERTICAL  SECTION. 

provided  with  cast-iron  lugs  for  tightening.  They  resist  chemical 
action  for  a  long  time,  while  flat  hoops  are  quickly  destroyed. 
It  is  well  to  paint  them  with  two  coats  of  asphaltum  varnish 
before  placing  them.  One  of  them  should  be  placed  at  a  level 
with  the  bottom,  or  only  a  very  little  below,  so  as  not  to  interfere 
with  the  outlet  pipe  of  the  tank.  The  hoops  should  be  12  in. 
apart.  The  bed  on  which  the  tank  rests  should  be  made  so  that 
the  tank  inclines  toward  the  front  1  to  1£  in. 

If  a  tank  has  been  in  operation  for  some  time  it  will  be  found 
that  some  material  has  accumulated  under  the  filter  bottom, 


LIXIVIATION   WITH  SODIUM  HYPOSULPHITE 


159 


which  material  has  to  be  removed  from  time  to  time,  so  as  to 
permit  a  full  flow  of  the  filtrate  from  all  parts  toward  the  outlet. 
To  facilitate  this  cleaning  the  filter  bottom  should  be  made  in 
sections,  small  enough  to  be  handled  by  two  men.  To  the  pieces 
A,  (Fig.  31),  which  are  4  in.  high  and  2i  in.  wide,  1  x  1  in. 
laths  are  nailed  1  in.  apart.  Square  notches  cut  into  the  pieces 


FIG.  32.  — LEACHING  TANK,  PLAN. 

This  shows  filter  bottom  arranged  in  fourteen  sections.    Two  sections  are  complete. 

A  permit  the  flow  of  the  solution  between  the  sections.  Fig.  32 
illustrates  the  arrangement  of  these  sections,  showing  two  of 
them  completed.  All  the  lumber  of  which  the  filter  bottoms 
are  made  should  be  planed  and  painted  with  asphaltum  varnish, 
and  the  heads  of  the  nails  well  driven  into  the  laths,  and  covered 
with  white  lead  and  asphaltum.  The  upper  edges  of  the  laths 


160  HYDROMETALLURGY   OF  SILVER 

should  be  beveled  to  prevent  them  from  cutting  the  filter  cloth. 
For  economical  reasons  it  is  well  to  order  the  laths  together  with 
the  tanks,  and  not  attempt  to  make  them  at  the  mine. 

About  one  inch  above  the  filter  bottom  the  tank  is  provided 
with  a  groove.  This  groove  is  made  by  nailing  first  a  wooden 
strip  1  x  2  in.  around  the  inner  side  of  the  tank,  then  by  nailing 
to  this  strip  another  1x4  in.  This  gives  a  groove  1x2  in. 
deep.  It  serves  to  tuck  in  the  ends  of  the  filter  cloth.  B,  B, 
Figs.  31  and  32,  are  the  lead  outlet  pipes.  It  is  not  advisable  to 
use  larger  outlet  pipes  than  \\  in.  because  they  have  to  be  con- 
nected with  a  rubber  hose  about  5  ft.  long,  and  if  they  are  of 
too  large  a  diameter,  flatten,  and  do  not  permit  a  flow  in  propor- 
tion to  their  size,  as  the  solution  has  no  pressure.  Over  a  \\-rn. 
lead  pipe  a  IJ-in.  hose  can  be  pushed,  and  if  it  is  5-ply  will  remain 
round  and  discharge  a  full  stream.  If  the  solution  flows  out 
with  force  it  is  a  sign  that  the  outlet  is  not  large  enough,  and  a 
second  or  even  a  third  one  has  to  be  provided.  To  check  the 
stream  is  a  poor  practice,  because  a  free  percolation  is  of  great 
advantage,  as  well  with  regard  to  the  extraction  as  to  the  time 
required  for  it. 

Under  but  close  to  the  top  of  the  filter  bottom  is  the  J-in. 
or  f-in.  lead  pipe  N ',  Fig.  31.  This  pipe  is  connected  with  a 
rubber  hose  which  extends  up  to  the  rim  of  the  tank.  The  hose 
can  be  closed  tight  by  a  'clamp-screw.  This  pipe  serves  to  permit 
the  air  from  under  the  filter  to  escape.  This  is  of  importance  if 
the  working  of  a  tank  has  to  be  interrupted.  If  the  solution 
outlet  is  stopped,  the  air  under  the  filter  will  be  brought  under 
pressure  by  the  solution,  and,  if  there  is  no  escape  for  it,  it  is  apt 
to  break  through  the  filter. 

I  have  found  the  use  of  hose  clamps  in  lixiviation  works  not 
only  very  convenient  but  also  economical.  Valves  are  more  or 
less  corroded  by  the  solutions  and  are  of  short  life,  while  the 
clamp,  doing  the  service  of  a  valve,  does  not  come  in  contact 
with  the  solutions.  These  clamps  are  best  when  made  of  brass. 
They  are  not  in  the  market  and  have  to  be  made  to  order.  Figs. 
33  and  34  give  the  construction  and  size  of  a  clamp  for  a  li  to 
2-in.  hose.  Smaller  sizes  for  smaller  hose  have  to  be  kept  on 
hand  also. 

The  tanks  ought  not  to  be  placed  on  a  solid  terrace,  as  we 
often  find  done;  it  is  much  more  rational  to  place  them  on 


LIXIVIATION   WITH   SODIUM   HYPOSULPHITE 


161 


trestles,  because  a  leak  in  the  tank  is  easily  discovered,  and, 
besides,  it  gives  room  underneath  to  run  cars  for  the  removal  of 
the  residues. 

Solubility  of  Silver  Chloride  in  Base-Metal  Chlorides  and  Salt.  — 
The  base-metal  chlorides  as  well  as  sodium  chloride  have  the 
property  of  dissolving  silver  chloride,  their  dissolving  energy 
increasing  with  the  temperature  and  concentration.  It  is  there- 


FIGS.  33  and  34.  — BRASS  CLAMPS  FOR  1J-AND  2-IN.  HOSE. 
Full  size. 

fore  not  advisable  to  leach  with  hot  water  or  to  use  too  deep 
leaching  vats,  even  if  the  filtering  quality  of  the  ore  permits  it, 
because  the  water  by  passing  through  a  thick  layer  of  ore  becomes 
highly  charged  with  these  salts  and  considerable  silver  chloride 
will  be  dissolved.  A  very  dilute  chloride  solution  does  not 
dissolve  silver  chloride,  and  in  order  to  prevent  the  base-metal 
chlorides  from  dissolving  silver,  the  washing  of  the  ore  should  be 


162  HYDROMETALLURGY  OF  SILVER 

so  conducted  that  a  sufficiently  dilute  solution  will  be  pro- 
duced. 

Prevention  of  Loss  of  Silver  in  Base-Metal  Solutions.  —  In 
tank  lixiviation  this  can  only  be  accomplished  to  a  certain  degree 
(not  perfect)  by  two  methods  which  were  introduced  by  myself: 
(1)  The  water  is  allowed  to  enter  the  vat  below  the  filter  until  it 
gradually  rises  above  the  ore.  Thus  the  most  concentrated 
portion  of  the  solution  will  appear  above  the  ore,  and  if  then 
diluted  with  a  stream  of  water  and  the  course  of  filtration  reversed 
a  large  part  of  the  silver  chloride  will  be  precipitated  and  remain 
in  the  ore,  while  the  outflowing  solution  will  be  dilute,  although 
not  sufficiently  so  as  to  be  entirely  free  of  silver.  (2)  A  quicker 
but  not  quite  so  effective  a  method  is  to  fill  the  tank  partly  with 
water  and  then  to  dump  the  ore  into  it,  either  dry  or  moist. 
Enough  water  should  be  taken  to  rise  about  4  in.  above  the 
charge  when  complete.  By  this  method  the  outflowing  solution 
will  not  be  so  dilute  as  by  the  first,  but  much  more  so  than  it 
would  be  if  the  ore  were  charged  in  an  empty  tank  and  then  leached 
from  above.  However,  the  only  method  by  which  this  problem 
is  actually  solved  is  trough  lixiviation,  also  introduced  by  me. 
By  this  operation  the  ore  can  be  brought  at  once  in  contact  with 
such  quantities  of  water  as  to  produce  a  solution  so  dilute  that 
not  a  trace  of  silver  will  be  dissolved.  More  about  this  method 
of  lixiviation  will  be  said  further  on. 

The  amount  of  silver  dissolved  during  base-metal  leaching 
varies  greatly,  and  depends  on  the  nature  of  the  ore,  the  thickness 
of  the  layer,  the  amount  of  salt  used  in  roasting,  the  mode  of 
treatment  in  the  vat,  and  the  temperature  of  the  outflowing 
solution.  In  some  works,  usually  in  large  ones,  the  ore  is  charged 
steaming  hot  from  the  cooling  floor,  or  if  for  special  reason  it  be 
charged  dry  it  will  enter  the  tank  even  hotter,  and  as  a  matter 
of  course  the  first  part  of  the  outflowing  solution  will  be  very 
hot  and  contain  considerable  silver.  To  illustrate  how  the 
amount  of  silver  dissolved  by  the  base-metal  solution  varies 
under  different  conditions,  a  few  examples  may  be  given: 

San  Francisco  del  Oro,  Parral,  Mexico.  —  Heavy  zinc-lead  ore 
with  iron  pyrites,  roasted  with  4  per  cent,  salt ;  the  tanks  partly 
filled  with  water  and  the  ore  dumped  dry  and  cool  into  it ;  average 
value  of  roasted  ore  26.1  oz.  per  ton;  charge,  8  tons;  leaching 
time,  8  hours;  silver  dissolved,  1  per  cent. 


LIXIVIATION   WITH   SODIUM  HYPOSULPHITE  163 

Sombrerete,  Zacatecas,  Mexico.  —  Heavy  ore  containing  zinc 
blende,  galena,  iron  pyrites  and  some  sulphureted  copper  minerals, 
roasted  with  6  per  cent,  salt;  the  tank  partly  filled  with  water 
and  the  ore  dumped  dry  and  warm  into  it;  average  value  of 
roasted  ore,  42.6  oz.  silver  per  ton;  charge,  52.5  tons;  leaching 
time,  12  hours;  silver  dissolved,  1.23  per  cent. 

Cusihuiriachic,  Chihuahua,  Mexico.  —  Not  so  heavy  an  ore, 
containing  galena,  zinc  blende,  some  iron  and  copper  pyrites, 
some  silver-copper  glance  and  also  some  ruby  silver;  roasted  with 
8  per  cent,  salt;  the  ore  wetted  on  the  cooling  floor  and  charged 
slightly  warm  into  an  empty  tank  and  leached  from  above; 
average  value  of  roasted  ore,  47.9  oz.  silver  per  ton;  charge, 
8  tons;  leaching  time,  53  hours;  silver  dissolved,  2.5  per  cent. 

Determination  of  Amount  of  Silver  Dissolved.  —  Mr.  Russell 
reports  the  amount  of  silver  dissolved  from  the  Cusihuiriachic  ore 
as  11.6  per  cent,  if  the  ore  is  charged  cold  and  dry  and  leached 
from  above.  Though  Mr.  Russell  roasted  with  10  per  cent,  of 
salt  I  believe  his  figure  to  be  much  too  high.  His  method  of 
ascertaining  the  amount  of  silver  dissolved  is  not  correct.  He 
determines  on  a  sample  in  the  laboratory  the  amount  of  salts 
soluble  in  water  contained  by  the  roasted  ore,  and  from  this 
figure  he  calculates  the  amount  of  silver  which  ought  to  be  con- 
tained in  the  ore  after  leaching  with  water,  compares  it  with  the 
actual  amount  found  in  it  by  assay,  and  then  estimates  the 
amount  of  silver  dissolved.  This  is  not  correct,  because  a  large 
charge  in  the  works  cannot  be  so  thoroughly  washed  as  a  small 
sample  in  the  laboratory,  and  therefore  the  calculated  value  of 
the  washed  ore  will  be  much  too  high,  and  consequently  the 
calculated  amount  of  silver  dissolved. 

The  correct  way  to  ascertain  the  amount  of  silver  dissolved  is 
to  collect  separately  in  vats  the  whole  base-metal  solution  of  one 
charge  and  measure  its  volume.  If  it  be  then  determined  how 
much  silver  is  contained  in  1000  c.c.,  it  is  an  easy  matter  to 
calculate  the  total  amount  of  silver  dissolved.  To  take  the 
sample  out  of  the  vats  would  not  be  correct,  because  the  solution 
being  collected  from  different  vats  will  not  be  uniform  in  concen- 
tration and  therefore  will  not  be  uniform  as  to  tenor  of  silver. 
Besides,  a  large  portion  of  the  silver  will,  be  precipitated  by  the 
gradual  dilution  of  the  first  concentrated  brine,  which,  to  a 
certain  extent,  will  escape  the  sample,  even  if  the  contents  of  the 


164  HYDROMETALLURGY   OF  SILVER 

vat  be  agitated.  However,  there  is  no  difficulty  in  obtaining  a 
correct  sample.  A  small  rubber  tube,  terminating  with  a  glass 
tube  drawn  to  a  fine  point,  is  inserted  in  the  outlet  of  the  vat  and 
left  there  during  the  whole  time  of  base-metal  leaching.  Thus  a 
very  fine  stream  of  the  outflowing  solution  is  obtained,  and  if 
collected  in  a  glass  vessel  of  proper  size,  a  representative  sample 
of  about  two  or  three  gallons  of  the  whole  solution  is  obtained. 
The  volume  of  this  sample  is  correctly  measured  and  all  the 
heavy  metal  salts  contained  therein  precipitated  with  calcium  or 
sodium  sulphide.  The  precipitate  is  collected  on  a  filter,  dried, 
weighed  and  assayed,  and  the  amount  of  silver  contained  in  it  is 
calculated.  With  this  figure  and  the  total  volume  of  solution 
collected  in  the  vats  the  total  amount  of  silver  dissolved  during 
base-metal  leaching  may  be  estimated,  and  by  knowing  the 
weight  of  the  charge  the  amount  dissolved  per  ton  can  be  calcu- 
lated and  expressed  in  percentage.  The  figures  of  silver  dissolved 
under  different  conditions  quoted  above  were  obtained  by  this 
method  and  therefore  may  be  considered  correct. 

Effect  of  an  Excess  of  Salt.  —  If  the  proper  amount  of  salt  be 
used  in  roasting,  so  that  all,  or  nearly  all,  be  decomposed,  and 
the  leaching  be  done  by  one  of  the  improved  methods,  it  may  be 
assumed  that  the  amount  of  silver  dissolved  will  never  exceed 
3  per  cent.  An  excess  of  salt  will  cause  a  large  percentage  of 
silver  to  be  dissolved,  reaching  60  to  70  per  cent,  in  my  own 
experience.  In  this  case  the  ore  was  of  such  a  nature  that  it 
required  a  large  excess  of  salt  in  roasting  to  produce  a  satisfactory 
chlorination  of  the  silver.  This  very  interesting  occurrence  will 
be  discussed  below. 

Precipitation  of  Silver  from  Base-Metal  Solutions.  —  There 
are  several  ways  of  precipitating  the  silver  contained  in  the  base- 
metal  solution,  including  precipitation  by  dilution  with  water, 
to  be  described  further  on,  and  precipitation  by  one  of  the  follow- 
ing reagents:  (1)  Milk  of  lime.  This  produces  a  very  voluminous 
precipitate  which  is  rather  difficult  to  handle,  and  from  which 
the  silver  cannot  be  extracted  unless  there  is  a  smelting  furnace 
connected  with  the  leaching  works.  In  such  a  case  lime  may 
be  used  to  advantage,  especially  since  it  is  so  cheap  a  reagent, 
but  care  must  be  taken  not  to  use  the  base-metal  solution  too 
concentrated,  in  which  case  it  is  coagulated  and  much  trouble 
in  further  handling  is  caused.  (2)  Calcium  or  sodium  sulphide. 


LIXIVIATION   WITH  SODIUM   HYPOSULPHITE  165 

These  precipitants  are  used  in  most  of  the  lixiviating  works,  but 
they  are  rather  expensive,  because  they  precipitate  base  metals 
also.  If  sodium  sulphide  be  added  gradually  to  a  tank  charge 
of  base-metal  solution,  it  will  be  observed  that  the  precipitate 
formed  first  is  the  richest  in  silver,  and  that  it  becomes  poorer  as 
precipitation  goes  on.  A  complete  precipitation,  however,  can  be 
produced  only  when  all  the  base  metals  are  thrown  down.  Since 
the  principal  portion  of  the  silver  is  precipitated  in  the  beginning, 
and  only  the  smaller  portion  remains  in  solution  with  the  main 
bulk  of  the  base-metal  salts,  the  point  will  be  reached  in  precipi- 
tating when  the  cost  of  the  precipitant  will  exceed  the  value  of 
the  precipitated  silver,  and  therefore  a  complete  precipitation 
is  not  advisable.  In  most  lixiviation  works  only  a  partial  pre- 
cipitation is  performed,  after  which  the  solution  with  the  re- 
mainder of  the  silver,  which  it  did  not  pay  to  precipitate,  is 
allowed  to  run  to  waste. 

Recovery  of  Silver  from  Waste  Liquor  by  Precipitation  with 
Copper.  —  Although  the  waste  liquor  will  be  found  very  low 
in  silver,  still  if  its  large  volume  be  taken  in  consideration  it  will 
appear  that  the  loss  occurring  in  this  way  during  a  year  is  im- 
portant. This  loss,  however,  can  be  diminished  greatly  by  pre- 
cipitation with  copper,  which  acts  most  energetically  if  finely 
divided  like  cement  copper,  and  if  the  solution  be  warm.  To 
effect  a  perfect  desilverization  it  is  necessary  to  treat  the  heated 
solution  in  charges  and  to  use  mechanical  appliances  to  produce 
an  intimate  contact  with  the  cement  copper,  but  the  base-metal 
solution,  after  a  partial  precipitation  with  calcium  or  sodium  sul- 
phide, is  not  rich  enough  to  warrant  an  expensive  and  complicated 
treatment,  and  it  will  be  found  to  be  more  profitable  to  employ 
a  more  primitive  method,  even  if  not  all  of  the  silver  is  recovered. 
The  true  economy  in  metallurgy,  as  in  any  other  industry,  is  to 
save  the  most  at  the  least  expense.  As  soon  as  its  expense  ex- 
ceeds the  value  of  the  recovered  metal,  a  method  ceases  to  be 
practicable,  no  matter  how  interesting  it  may  be.  That  method 
should  be  adopted  which  is  best  suited  to  local  conditions. 

The  largest  part  of  the  dissolved  silver  can  be  saved  by  con- 
veying the  solution  through  a  series  of  flat  tanks  in  which  cement 
copper  is  so  divided  that  it  offers  a  large  surface  of  contact  to 
the  solution.  The  exhaust  steam  of  the  engine  can  be  used  to 
increase  the  temperature  of  the  stream,  the  steam  being  made 


166  HYDROMETALLURGY  OF  SILVER 

to  enter  the  solution  through  pipes  about  12  to  18  in.  below  the 
surface  in  different  tanks  of  the  system.  If  the  base-metal  solu- 
tion contains  copper,  the  pipe  projecting  into  the  solution  ought 
to  be  of  lead.  The  steam  being  condensed  a  vacuum  is  formed, 
and  very  little  if  any  back  pressure  to  the  engine  will  be  noticed. 
Wherever  the  climate  allows  it  these  tanks  can  be  built  in  the 
yard  outside  the  works,  without  roof  or  shelter.  If  the  ore  con- 
tains copper,  sufficient  cement  copper  will  be  formed  by  placing 
scrap  iron  in  the  tanks,  but  if  copper  is  wanting,  the  cement 
copper  has  to  be  made  from  a  dilute  solution  of  blue  vitriol. 
Ores  which  are  treated  by  lixiviation  are  seldom  entirely  free 
from  copper,  and  for  this  reason  it  is  well  to  place  some  scrap 
iron  with  the  cement  copper. 

Base-Metal  Leaching  at  Sombrerete.  —  At  Sombrerete,  Mexico, 
I  conducted  the  base-metal  leaching  in  the  following  manner: 
The  ore  is  very  permeable,  and  Stetefeldt  and  Russell,  who  built  the 
works,  erected  leaching  vats  15  ft.  6  in.  in  diameter  and  7  ft.  6  in. 
deep,  capable  of  receiving  55  to  58  tons  of  roasted  ore.  These 
tanks  were  decidedly  too  deep.  The  ore  filtered  freely  enough  to 
permit  a  deep  charge,  and  in  silver  leaching  this  would  not  inter- 
fere, but  in  base-metal  leaching  it  caused  much  trouble.  The 
ore  was  not  wetted  on  the  cooling  floor  to  produce  additional 
chlorination,  but  was  charged  dry.  The  water  passing  through  7 
ft.  of  roasted  ore  became  so  saturated  with  salts  that  they  crystal- 
lized and  blocked  the  filter,  the  space  below  the  filter,  and  the 
outlet  pipe,  which  interfered  greatly  with  the  work  and  was  very 
annoying.  I  overcame  this  difficulty  by  filling  the  vats  to  the 
depth  of  about  3  ft.  with  water  and  dumping  the  ore  into  it;  fre- 
quently the  ore  had  to  be  charged  while  quite  hot  in  order  to 
make  room  on  the  cooling  floor,  and  consequently  the  outflowing 
base-metal  solution  was  warm  and  dissolved  more  silver  chloride 
than  it  would  have  done  if  more  favorable  conditions  could  have 
been  maintained.  But  notwithstanding  this,  the  amount  of 
silver  dissolved  was  much  less  than  when  the  ore  was  leached 
from  above. 

If  a  base-metal  solution  containing  silver  chloride  be  diluted 
with  water  the  silver  chloride  will  precipitate  as  such.  In  the 
Sombrerete  works  there  were  two  base-metal  precipitation  vats 
9  ft.  9  in.  in  diameter  and  9  ft.  deep;  and  in  order  to  take  advan- 
tage of  the  above  reaction  I  allowed  the  base-metal  solution  to 


LIXIVIATION   WITH   SODIUM  HYPOSULPHITE  167 

run  into  both  simultaneously,  by  which  method  I  got  the  con- 
centrated portion  of  the  solution,  which  contained  the  most  silver, 
divided  evenly  into  the  two  vats.  This  left  considerable  room 
in  them  for  diluting,  and  by  allowing  the  less  concentrated  and, 
finally,  the  very  weak  solution  to  run  evenly  into  both  vats,  I 
obtained  a  uniform  solution  dilute  enough  to  cause  the  main 
portion  of  the  dissolved  silver  chloride  to  be  precipitated.  Silver 
chloride  thus  precipitated  being  very  finely  divided  requires  a 
long  time  to  settle.  For  the  purpose  of  effecting  quick  settling 
and  at  the  same  time  precipitating  more  silver,  5  to  10  gal.  of 
sodium  sulphide  solution  were  added  to  each  vat  and  agitated. 
In  this  way  a  part  of  the  silver  chloride  was  converted  into  sul- 
phide, while  the  undecomposed  part  was  readily  collected  by  the 
flocculent  precipitate  of  copper,  lead  and  other  base  metals,  and 
after  agitation  was  interrupted  the  whole  precipitate  would 
settle  quickly. 

After  the  precipitate  has  settled,  the  clear  solution  is  decanted 
by  a  stiff  rubber  hose,  which  enters  the  vat  close  to  the  bottom 
and  projects  above  the  surface  of  the  liquor.  By  lowering  this  a 
few  inches  below  the  surface  the  liquor  flows  out,  and  if  the  hose 
is  lowered  gradually  as  the  liquor  runs  out  it  is  always  the  clearest 
part  of  the  latter  which  is  drawn  off.  This  method  of  decanta- 
tion  is  better  than  a  series  of  tubes  inserted  at  different  levels  in 
the  side  of  the  vat,  because  during  precipitation  these  tubes  are 
filled  with  precipitate,  and  in  decanting  the  first  solution  from 
each  tube  contains  precipitate  and  consequently  has  to  be  con- 
veyed to  filters. 

Outside  the  building  were  constructed  a  series  of  flat  square 
tanks  2J  ft.  deep,  built  of  stone  and  mortar,  well  plastered  and 
coated  with  asphaltum  varnish,  which  were  provided  with  a 
number  of  movable  wooden  double  benches,  or  shelves,  loaded 
down  with  scrap  iron.  In  this  way  the  scrap  iron  was  well  dis- 
tributed throughout  the  tanks  and  offered  a  large  surface  to  the 
solution.  They  were  so  arranged  that  the  flowing  solution 
would  move  in  its  whole  depth  and  had  to  take  a  zigzag  course. 

The  ore  of  Sombrerete  contains  about  2  per  cent,  copper,  and 
the  outflowing  solution  for  a  considerable  time  is  colored  green 
by  cupric  chloride,  and  a  good  deal  of  copper  is  precipitated  by 
the  scrap  iron.  The  precipitated  cement  copper  incrusting  the 
iron  is  loose  and  spongy  and  offers  a  very  large  surface  for  pre- 


168    '  HYDROMETALLURGY  OF  SILVER 

cipitating  the  silver.  It  ought  not  to  be  disturbed  by  stirring 
the  solution,  since  thereby  the  copper  falls  to  the  bottom  and 
does  much  less  service.  However,  it  is  difficult  to  precipitate 
all  the  silver,  and  it  requires  a  rather  large  number  of  tanks. 
Silver  dissolved  in  a  cupric  chloride  solution  cannot  be  completely 
precipitated  until  all  the  cupric  chloride  is  decomposed;  there- 
fore as  long  as  the  solution  leaving  the  last  tank  still  gives  a 
reaction  for  copper  it  can  be  assumed  to  contain  some  silver  also. 
This  indicates  the  necessity  of  erecting  one  or  more  additional 
tanks.  In  Sombrerete  these  tanks  were  cleaned  once  a  month, 
and  the  cement  copper  obtained  from  them  contained  500  to  600 
oz.  silver  per  ton,  and  60  to  70  per  cent,  copper. 

Cupric  chloride  in  solution  in  contact  with  cement  copper, 
especially  at  an  elevated  temperature,  is  converted  into  cuprous 
chloride,  which  settles  as  a  heavy  white  crystalline  precipitate. 
Cuprous  chloride  again  in  contact  with  metallic  iron  is  converted 
into  metallic  copper,  and  the  iron  into  ferrous  chloride.  It  is 
clear  that  in  these  tanks,  in  which  the  cement  copper  surrounds 
the  iron,  an  intimate  contact  between  the  iron  and  the  cuprous 
chloride  is  not  possible,  and  that  consequently  the  cement  copper 
will  contain  a  large  percentage  of  cuprous  chloride.  This  is  not  a 
desirable  associate  for  the  cement  copper,  because  in  the  subse- 
quent treatment  it  always  causes  a  large  loss  of  copper  and  silver, 
and  therefore  the  cuprous  chloride  ought  to  be  decomposed. 
This  should  be  done  shortly  after  the  cement  copper  is  taken 
out  of  the  tanks,  because  if  left  exposed  to  the  air  the  cuprous 
chloride  will  change  into  oxichloride.  The  quickest  and  most 
rational  way  is  to  charge  the  cement  copper  in  a  revolving  barrel 
with  an  addition  of  water,  salt  and  some  light  scrap  iron.  The 
barrel  is  made  of  3-in.  staves  and  heads,  and  is  bound  with  copper 
hoops.  It  is  constructed  like  an  amalgamation  barrel.  The 
inside  is  lined  with  hardwood  lath  about  1J  in.  thick,  so  that  the 
body  of  the  barrel  is  protected  from  the  wear  and  tear.  Four 
longitudinal  ribs,  projecting  about  2J  in.,  are  inserted  diametri- 
cally in  the  lining,  which  produces  a  lively  mixing  of  the  charge. 
The  axles,  or  trunnions,  which  are  made  of  brass,  or  better  of 
bronze,  are  provided  with  a  strong  flange  by  which  they  are 
bolted  to  the  heads  of  the  barrel.  Each  is  bored  and  provided 
with  a  stuffing-box  through  which  a  copper  pipe  enters.  One  of 
the  pipes  is  turned  downward  close  to  where  it  enters,  but  must 


LIXIVIATION   WITH  SODIUM  HYPOSULPHITE  169 

not  project  so  far  into  the  lower  half  of  the  barrel  that  the  scrap 
iron  and  the  cement  copper  will  strike  it.  Through  this  pipe 
steam  is  introduced  to  heat  the  pulp.  The  pipe  that  passes 
through  the  other  trunnion  is  turned  upward  high  enough  to 
just  clear  the  ribs  of  the  barrel.  This  pipe  serves  as  outlet  for 
the  gases  and  steam  and  leads  outside  the  building.  A  manhole 
on  the  side  of  the  barrel,  provided  with  a  copper  frame  and  lid, 
serves  for  charging  and  discharging,  while  another  in  the  head 
gives  entrance  in  case  the  lining  is  to  be  renewed  or  other  re- 
pairs are  to  be  made. 

When  the  barrel  is  charged  with  cement  copper,  water  and 
iron,  salt  is  added,  steam  turned  on,  and  the  barrel  is  put  in 
revolution.  A  warm  solution  of  salt  (sodium  chloride)  dissolves 
readily  cuprous  chloride,  and  from  this  solution  the  iron  precipi- 
tates the  copper  with  great  energy.  Water  should  be  used 
moderately,  just  enough  to  produce  a  lively  movement  of  the  pulp, 
say  three  to  four  times  the  volume  of  the  cement  copper,  and  to 
produce  a  strong  brine  without  using  too  much  salt.  The  reaction 
creates  heat  and  the  escape  pipe  has  to  be  watched.  When  the 
steam  commences  to  come  out  freely  the  steam  inlet  ought  to  be 
closed.  The  side  of  the  barrel  is  provided  with  a  plug-hole 
through  which  the  operator  can  take  a  sample  by  means  of  a 
flask  fastened  to  a  copper  wire.  If  the  filtrate  does  not  show  a 
reaction  for  copper  the  process  is  finished.  This  takes  about  45 
minutes.  Below  the  barrel  is  placed  a  square  flat  filter  tank, 
above  which  is  a  large,  strong  iron  screen  with  2J-in.  holes,  the 
frame  of  which  rests  on  four  car-wheels  on  rails,  so  that  it  can  be 
moved  easily  or  withdrawn  entirely  from  the  tank.  The  con- 
tents of  the  barrel  are  discharged  on  this  screen,  and  the  cement 
copper  washed  into  the  filter  tank  by  a  stream  of  water,  while 
the  scrap  iron  remains  on  the  screen.  Then  some  sulphuric  acid 
is  added  to  remove  the  basic  iron.  This  acid  solution  is  allowed 
to  pass  through  the  cement  copper,  after  which  the  latter  is  well 
washed.  By  this  method  I  obtained  cement  copper  containing 
90  to  95  per  cent,  copper. 

If  the  cement  copper  has  to  be  prepared  from  blue  vitriol  it 
may  be  desirable  to  use  it  for  a  longer  time,  in  which  case  it  will 
become  much  richer  in  silver,  but  it  has  to  be  taken  out  occa- 
sionally and  treated  with  dilute  sulphuric  acid  to  remove  the  basic 
salts.  Thus  purified  the  cement  copper  acts  again  with  energy. 


170 


HYDROMETALLURGY  OF  SILVER 


In  order  not  to  interfere  with  the  regular  work,  the  tanks 
have  to  be  so  arranged  that  during  cleaning  half  of  them  remain 
in  operation  while  the  others  are  disconnected  and  cleaned.  As 
soon  as  half  are  cleaned,  the  scrap  iron  or  cement  copper  is  put 
in  place  and  the  solution  allowed  to  enter  again  while  the  other 
tanks  are  cleaned. 

Precipitating  the  Dissolved  Silver  Chloride  by  Dilution  with 
Water.  —  This  method  was  first  recommended  and  introduced  by 
me.  All  alkaline  and  metal  chlorides  have  the  property,  when 
concentrated,  of  dissolving  silver  chloride,  and  dropping  it  again 
as  such  when  diluted  wTith  water.  The  precipitation  takes  place 
in  proportion  to  the  dilution,  and  if  sufficient  water  be  added  all 
the  silver  will  be  precipitated.  This  method  of  desilverizing  the 
base-metal  solution  is  undoubtedly  the  most  effective  and  cheapest, 
all  that  is  required  being  a  sufficient  supply  of  water  and  a  few 
more  vats  for  base-metal  precipitation  than  are  usually  found  in 
lixiviation  works.  I  used  this  method  first  at  the  Silver  King 
mill  in  Arizona  in  1880  to  1882  with  very  good  results,  then  in 
various  other  localities  where  the  supply  of  water  permitted  it, 
and  also  at  the  mill  of  the  Hidalgo  Mining  Company,  at  Parral, 
Mexico,  in  1894. 

Experience  at  Parral,  Mexico.  —  The  observations  I  made  at 
Parral  were  very  interesting,  and  I  shall  give  them  in  detail. 
The  ore  which  was  treated  consisted  principally  of  galena,  lead 
carbonate,  and  blende,  and  was  almost  free  from  iron  pyrites. 
Neither  galena  nor  blende  produces  in  roasting  with  salt  much 
chlorine,  especially  if  the  blende  belongs  to  that  variety  which 
contains  little  or  no  iron  sulphide.  In  this  case  the  chlorination 
of  the  silver  depended  principally  on  the  chlorine  produced  by 
the  action  of  the  quartz  on  the  salt  and  by  the  direct  action  of 
volatilized  sodium  chloride.  Such  roasting  requires  a  large  excess 
of  salt  and  a  high  heat.  The  roasting  was  conducted  in  a  White- 
Howell  furnace  with  9  to  10  per  cent,  of  salt.  Experiments  were 
made  with  the  aim  of  reducing  the  proportion  of  salt,  but  as  soon 
as  the  amount  fell  below  9  per  cent,  the  extraction  suffered  so 
much  that  it  was  more  rational  to  use  a  higher  percentage  of 
salt.  If  the  ore  had  contained  sufficient  iron  pyrites  3  to  4  per 
cent,  of  salt  would  have  been  sufficient,  because  lead-zinc  ores  in 
presence  of  iron  pyrites  require  less  salt  for  a  successful  roasting 
than  any  other  class  of  complex  ores;  but  since  the  iron  pyrites 


LIXIVIATION   WITH  SODIUM  HYPOSULPHITE  171 

was  wanting  and  an  excess  of  salt  had  to  be  used,  and  since  zinc 
blende  and  galena  act  but  slightly  on  the  salt,  the  roasted  ore 
contained  a  large  amount  of  sodium  chloride. 

The  lixiviation  mill  of  the  Hidalgo  Mining  Company  is  well 
constructed  and  arranged,  and  reflects  credit  on  J.  T.  Long,  who 
erected  it.  This  mill  has  a  very  large  cooling  floor,  on  which  the 
roasted  ore  is  allowed  to  cool  dry  for  three  days.  Notwithstanding 
this  long  time,  the  ore,  when  charged,  is  still  hot  enough  to  impart 
to  the  outflowing  base-metal  solution  for  a  rather  long  time'  a 
temperature  of  140  to  200  deg.  F.  The  excess  of  salt  contained 
in  the  roasted  ore  dissolved  readily  in  the  water  of  this  temper- 
ature, forming  a  highly  concentrated  brine  in  which  60  to  70 
per  cent,  of  the  silver  was  dissolved,  and  therefore  the  principal 
extraction  was  done  during  base-metal  leaching.  This,  of  course, 
required  a  careful  treatment  of  the  base-metal  solution. 

Sodium  sulphide  was  used  as  precipitant,  but  a  large  quantity 
was  required  on  account  of  the  large  amount  of  lead  contained 
in  the  solution.  If  a  zinc  blende  ore  containing  iron  pyrites  is 
roasted  with  the  proper  amount  of  salt  and  charged  cool  into  the 
vat  the  base-metal  solution  will  contain  but  little  lead,  if  any, 
provided  cold  water  is  used;  but  if  a  concentrated  hot  solution 
of  sodium  chloride  is  formed  in  leaching,  the  result  is  entirely 
different,  since  this  dissolves  not  only  lead  chloride,  but  also  lead 
sulphate,  and  the  solution  will  contain  large  quantities  of  these 
lead  salts.  If  the  solution  be  very  concentrated,  and  allowed  to 
cool,  large  crystals  will  be  formed,  while  the  mother  liquor  re- 
mains clear.  If  the  solution  be  not  very  concentrated  there  will 
be  only  a  crystalline  precipitate  which  turns  the  solution  milky. 
Neither  the  crystals  nor  the  crystalline  precipitate  contain  much 
silver.  The  main  part  of  the  silver  chloride,  lead  sulphate  and 
lead  chloride  remains  in  the  solution,  but  if  this  be  diluted  with 
water  it  turns  milky,  forming  a  heavy  white  precipitate;  and  if 
sufficient  water  is  used  the  white  precipitate  will  contain  all  the 
silver  and  lead  dissolved  in  the  solution  before  diluting. 

For  illustration  I  shall  record  my  observations  made  on  a 
certain  charge.  The  ore  when  charged  was  rather  hot.  The 
leaching  was  done  from  above.  The  solution  flowing  out  first 
had  a  temperature  of  200  deg.  F.  and  was  unusually  concentrated. 
Some  solution  was  collected  in  a  large  beaker  and  left  to  cool. 
A  large  amount  of  transparent  crystals  was  formed.  After  pouring 


172  HYDROMETALLURGY  OF  SILVER 

off  the  clear  mother  liquor  a  quantity  of  water  was  added  to  the 
latter,  which  caused  a  heavy  white  precipitate.  This  was  sepa- 
rated from  the  solution  by  filtration,  and  the  three  substances 
were  assayed  for  silver  with  the  following  results:  (1)  The  crystals 
contained  7.6  oz.  silver  per  ton;  (2)  the  white  precipitate  had 
3386  oz.  silver  per  ton;  (3)  the  filtrate  had  none.  Of  this  charge 
of  ore,  which  contained  19.6  oz.  silver  per  ton,  62.6  per  cent,  of 
the  silver  was  extracted  during  base-metal  leaching. 

Working  by  this  method  I  found  it  more  convenient  to  do  the 
precipitation  in  the  trough  leading  from  the  lixiviating  vats  to 
precipitation  vats,  by  allowing  a  stream  of  water  to  enter  the 
trough.  In  flowing  some  distance  the  solution  and  water  become 
thoroughly  mixed,  and  the  solution  is  perfectly  desilverized  before 
it  enters  the  vats.  The  stream  carries  along  the  white  precipitate, 
which  settles  in  the  precipitation  vats.  To  ascertain  whether  the 
solution  receives  the  proper  quantity  of  water  or  not,  a  sample 
of  the  diluted  solution  just  before  it  enters  the  vat  is  taken  and 
filtered.  Of  the  filtrate  50  c.c.  are  poured  into  a  beaker  and 
150  c.c.  of  cool,  clean  water  are  added.  If  it  remains  clear  after  a 
few  minutes  it  shows  that  no  more  water  is  needed;  if  it  turns 
milky  the  stream  of  water  has  to  be  increased.  Thus  by  decreas- 
ing or  increasing  the  stream  the  proper  dilution  can  be  main- 
tained. Of  course  the  base-metal  solution  can  also  be  diluted  in 
the  precipitation  vats,  but  by  this  method,  if  the  two  streams 
simultaneously  enter  the  vat,  it  is  difficult  to  ascertain  the  proper 
proportion,  and  if  part  of  the  vat  is  filled  with  solution  and  then 
water  is  added  there  will  be  a  considerable  loss  of  time.  Where 
mechanical  devices  are  used  for  agitation  it  is  advisable  to  have 
the  solution  agitated  while  the  vat  is  filling;  where  agitation  is 
performed  by  hand,  it  ought  to  be  done  from  time  to  time,  to 
make  the  precipitate  settle  more  quickly.  The  tanks  are  filled 
to  about  12  to  18  in.  below  the  rim,  to  leave  room  for  more  water 
in  case  the  proper  proportion  was  not  maintained  in  the  trough 
and  a  correction  is  necessary.  When  this  is  done  a  short  time  is 
allowed  for  the  heavy  part  of  the  precipitate  to  settle,  then  one 
quart  of  sodium  sulphide  is  added  and  the  solution  agitated 
again.  This  is  done  to  convert  the  very  fine  particles  of  the 
precipitate,  which  otherwise  would  remain  suspended  for  a  long 
time,  into  sulphide,  in  which  state  they  assume  a  flaky  condition 
and  settle  much  more  quickly. 


LIXIVIATION   WITH  SODIUM   HYPOSULPHITE  173 

The  lixiviation  tanks  were  charged  with  30  tons  of  roasted 
ore,  and  the  base-metal  leaching  required  about  twenty-four  hours. 
If  the  precipitation  was  not  done  by  water  but  by  sodium  sulphide, 
the  base-metal  solution  filled  7J-  tanks  of  3500  gal.  each,  and 
53  gal.  of  the  precipitant  were  used.  The  solution  running  to 
waste  still  contained  0.19  oz.  silver  per  1000  gal.,  or  0.15  oz.  per 
ton  of  ore  leached.  If  the  precipitation  was  carried  so  far  that 
all  the  base-metals  were  precipitated,  the  solution  running  to 
waste  did  not  contain  any  silver,  but  then  the  cost  of  the  precipi- 
tant exceeded  the  value  of  the  silver  saved.  If  the  precipitation 
was  effected  by  dilution  with  water  in  the  trough,  22  precipitation 
charges  of  3500  gal.  each  were  filled.  The  consumption  of  sodium 
sulphide  amounted  to  2.75  gal.  only,  and  the  clear  solution 
running  to  waste  did  not  contain  any  silver  at  all.  This  result 
shows  that  it  requires  on  an  average  about  two  parts  of  water  to 
one  part  of  solution.  Of  course,  in  the  beginning,  when  the  out- 
flowing solution  is  very  concentrated,  considerably  more  water 
has  to  be  added  than  in  the  above  proportion,  but  the  volume  of 
water  to  be  added  decreases  as  leaching  progresses,  and  toward 
the  end  it  is  less  than  the  volume  of  solution. 

On  account  of  the  large  quantity  of  lead  which  the  base-metal 
solution  carried,  and  which  is  precipitated  by  water  as  well  as  by 
sodium  sulphide,  there  was  not  much  difference  in  the  silver 
contents  of  the  precipitate  whether  obtained  by  water  or  by 
sodium  sulphide.  In  this  particular  case,  however,  a  great  differ- 
ence in  the  silver  contents  of  the  precipitate  was  observed  if  the 
water  for  leaching  was  first  applied  below  the  filter  or  from  the 
top.  In  the  former  instance  it  contained  1200  oz.  silver  per  ton; 
in  the  latter  only  800  oz. 

The  precipitation  with  water  is  undoubtedly  the  cheapest  and 
most  effective  mode  of  treating  the  base-metal  solution,  and  it 
ought  to  be  used  in  all  works  where  there  is  an  ample  supply. 

The  Use  of  Cupric  Chloride  During  Base-metal  Leaching.  — 
The  chlorination  of  the  silver,  and  correspondingly  the  extraction, 
can  be  increased  greatly  in  ores  which  do  not  contain  any  or  only 
a  little  copper,  by  use  of  a  solution  of  cupric  chloride  while  base- 
metal  leaching  is  in  operation.  The  cupric  chloride  can  be  pre- 
pared in  the  works  by  boiling  one  part  of  blue  vitriol  with  two 
parts  of  salt  in  a  wooden  vat  by  direct  application  of  steam. 
The  cupric  chloride  is  made  in  large  quantities  and  used  as  re- 


174  HYDROMETALLURGY   OF  SILVER 

quired.  It  takes  from  3  to  4  Ib.  of  blue  vitriol  and  6  to  8  Ib.  of 
salt  per  ton  of  ore.  To  avoid  the  formation  of  too  concentrated 
a  solution  the  cupric  chloride  is  not  added  until  the  leaching  has 
proceeded  one  or  two  hours.  It  is  better  not  to  add  the  whole 
of  it  at  once,  but  to  divide  it  so  that  one-fourth  is  added  to  about 
8  in.  of  water  above  the  ore;  and  when  this  charge  is  disappearing 
below  the  surface  another  charge  of  8  in.  water  and  the  second 
fourth  of  the  cupric  chloride  is  made,  and  so  on  until  all  is  used. 
If  the  outflowing  solution  is  green,  it  ought  to  be  collected  in  a 
separate  vat,  and  can  be  used  over  again  either  for  the  same 
charge  or  for  the  next  one  following,  according  to  circumstances. 
It  is  best  to  lift  this  solution  by  means  of  a  lead  steam  siphon  or 
injector.  If  an  ore  contains  so  much  copper  that  the  base-metal 
solution  is  colored  green,  this  reaction,  which  produces  an  addi- 
tional chlorination  of  the  silver,  takes  place  in  every  charge  in 
the  regular  operations  of  the  process,  and  this  is  one  of  the  reasons 
why  cupriferous  ores  are  the  easiest  to  treat  and  permit  the 
closest  extraction.  However,  even  with  such  ores,  better  results 
are  obtained  if  the  green  base-metal  solution  is  collected  separately 
and  allowed  to  pass  through  the  charge  a  second  time.  Any 
defect  in  the  roasting  will  be  much  lessened,  and  uniformly  good 
results  will  be  obtained.  The  more  inferior  the  roasting  the  higher 
will  be  the  percentage  of  gain  in  chlorination.  At  Parral,  where 
I  treated  the  San  Francisco  del  Oro  ore  containing  25.5  per  cent, 
zinc,  11.56  percent,  lead  and  1.02  per  cent,  copper,  on  imperfectly 
roasted  charges  by  the  use  of  cupric  chloride  32  to  38  per  cent, 
in  chlorination  were  gained.  At  the  Silver  King,  Arizona,  where 
I  treated  an  ore  containing  a  large  percentage  of  copper,  I  in- 
creased the  extraction  a  little  by  using  part  of  the  green  solution 
over  again.  With  ore  from  the  Veta  Grande  of  Parral,  which 
contained  mostly  zinc  blende,  galena,  and  lead  carbonate,  and 
no  copper,  I  experimented  on  a  large  scale  with  cupric  chloride, 
and  obtained  much  better  and  more  uniform  results. 

SILVER  LEACHING 

After  the  ore  is  freed  from  base-metal  salts  by  leaching  with 
water,  a  weak  solution  of  sodium  hyposulphite  is  introduced  on 
top  of  the  ore.  It  is  of  great  importance  to  work  with  a  weak 
solution.  If  the  solution  is  strong,  1,  2,  or  3  per  cent.,  and  an 
attempt  is  made  to  maintain  this  as  standard  strength,  it  can  be 


. 
LIXIVIATION  WITH  SODIUM  HYPOSULPHITE  175 

done  only  by  frequent  additions  of  sodium  hyposulphite;  else  the 
volume  of  the  stock  solution  will  decrease  rapidly  and  the  prepa- 
ration of  a  new  stock  solution  will  soon  be  required,  because  the 
ore  absorbs  a  large  volume  of  solution,  and  when,  after  the  silver 
extraction  is  concluded,  this  solution  is  displaced  by  water,  it  is 
impossible  to  regain  the  same  volume  of  the  original  strength. 
It  will  be  found  that  only  a  comparatively  small  portion  of  the 
outflowing  solution  will  have  the  standard  strength,  and  if  the 
displacement  of  the  solution  is  done  so  as  to  maintain  the  same 
volume  of  stock  solution,  much  water  will  enter  the  latter,  while 
considerable  hyposulphite  salt  will  remain  in  the  residues.  Such 
a  dilution  of  the  stock  solution  takes  place  also  in  the  beginning 
of  the  silver  leaching.  The  "  hypo  "  solution  passing  through  the 
washed  ore  has  to  displace  the  water  absorbed  by  the  ore,  and 
this  cannot  be  accomplished  without  the  first  portion  of  the 
solution  becoming  mixed  with  the  water;  and  since  even  a  very 
dilute  solution  of  sodium  hyposulphite  dissolves  silver  chloride, 
the  outflowing  solution  has  to  be  conveyed  to  the  silver  precipi- 
tation vats  a  long  time  before  it  reaches  its  standard  strength. 

Therefore  it  will  clearly  be  seen  that  the  standard  strength 
can  only  be  maintained  by  a  continual  addition  of  sodium  hypo- 
sulphite to  the  stock  solution.  A  continuous  supply  is  furnished 
by  the  precipitant,  whether  that  be  sodium  or  calcium  sulphide, 
because  in  preparing  it  some  sodium  or  calcium  hyposulphite  is 
formed.  As  much  as  5  to  7  per  cent,  of  this  salt  may  be  found 
in  the  freshly  prepared  precipitant,  without  the  latter  being 
exposed  to  the  action  of  the  air  for  any  considerable  length  of 
time.  The  constant  supply  from  this  source,  however,  does  not 
suffice  to  replace  the  loss  of  hyposulphite  salt  which  the  stock 
solution  suffers  if  the  latter  has  to  be  kept  at  a  strength  of  1  or 
2  per  cent.,  but  it  does  suffice,  and  in  fact  in  most  cases  exceeds 
the  loss,  if  a  weak  solution  is  used.  The  consequence  is  that  in 
the  first  case  sodium  hyposulphite  has  to  be  bought  and  added, 
which  increases  the  treatment  expense,  while  in  the  second  case 
not  a  pound  of  this  salt  has  to  be  added,  even  during  years  of 
operation,  and,  on  the  contrary,  water  has  often  to  be  added  to 
reduce  the  strength  to  the  standard. 

The  following  table  is  an  interesting  daily  record  of  the  sodium 
hyposulphite  contained  in  the  stock  solution  with  which  I  worked 
the  San  Francisco  del  Oro  ore.  With  the  exception  of  June  16, 


176 


HYDROMETALLURGY   OF   SILVER 


on  which  day,  for  a  certain  experimental  purpose,  175  Ib.  of 
sodium  hyposulphite  were  added,  there  was  not  a  pound  of  this 
salt  added  to  the  stock  solution: 


2 

2 

2 

2 

Kz 

K  z 

Kz 

Kz 

r.   O 

fe  O 

"•9 

OH 

0  H 

0  H 

OH 

] 

)ATE 

H  2 

DATE 

.  D 

] 

)ATE 

H  J 

] 

3ATE 

H§ 

we/) 

il 

Z  O 
WC/3 

If 

(6  ** 

OS  "" 

C6  " 

M  "* 

OH 

w 

0< 

£ 

£ 

AF 

>ril  13 

0.34 

May  11 

0.54 

Ju 

ne  7 

0.39 

Ju 

ly  4 

0.76 

14 

0.52 

"   12 

0.56 

8 

0.39 

5 

0.70 

15 

0.50 

"   13 

0.56 

9 

0.39 

6 

0.69 

16 

0.55 

"   14 

0.56 

10 

0.40 

7 

0.73 

17 

0.60 

"   15 

0.52 

11 

0.40 

8 

0.78 

18 

0.58 

"   16 

0.49 

12 

0.38 

9 

0.78 

19 

0.56 

"   17 

0.53 

13 

0.38 

10 

0.79' 

20 

0.61 

"   18 

0.53 

14 

0.38 

11 

0.81 

21 

0.60 

"   19 

0.52 

15 

0.38 

12 

0.63 

22 

0.58 

"   20 

0.57 

16 

0.50 

13 

0.61 

23 

0.58 

'   21 

0.57 

17 

0.50 

14 

0.58 

24 

0.58 

'   22 

0.59 

18 

0.49 

15 

0.55 

25 

0.57 

'   23 

0.58 

19 

0.48 

16 

0.56 

26 

0.56 

'   24 

0.54 

20 

0.53 

17 

0.59 

28 

0.53 

'   25 

0.56 

21 

0.55 

18 

0.60 

29 

0.54 

'   26 

0.58 

22 

0.55 

19 

0.62 

30 

0.51 

'   27 

0.60 

23 

0.56 

20 

0.61 

M 

y  1 

0.51 

'   28 

0.60 

24 

0.58 

21 

0.63 

2 

0.49 

'   29 

0.63 

25 

0.65 

22 

0.64 

3 

0.51 

'   30 

0.63 

26 

0-67 

23 

0.66 

4 

0.49 

'   31 

0.59 

27 

0.70 

24 

0.66 

5 

0.50 

June  1 

0.49 

28 

0.75 

25 

0.65 

6 

0.46 

"   2 

0.53 

29 

0.75 

26 

0.50 

7 

0.47 

"   3 

0.53 

30 

0.73 

27 

0.63 

8 

0.47 

"   4 

0.54 

Ju 

y  i 

0.80 

28 

(?) 

9 

0.50 

"   5 

0.48 

2 

0.81 

29 

0.59 

10 

0.49 

"   6 

0.43 

3 

0.79 

It  will  be  noticed  that  the  solution  had  a  tendency  to  increase 
in  strength.  In  order  to  keep  it  as  much  as  possible  at  the  stand- 
ard strength  of  0.50  per  cent,  water  had  to  be  added  frequently, 
which  accounts  for  most  of  the  sudden  drops.  A  sample  of  the 
stock  solution  taken  after  precipitation  contained:  hyposulphite 
salts,  0.52  per  cent.;  lime,  0.165  per  cent.;  sulphuric  acid,  0.140 
per  cent.;  chlorine,  0.098  per  cent.  The  solution  evaporated  and 
heated  in  the  muffle  gave  0.775  per  cent,  solids.  A  month  after, 
it  gave  0.785  per  cent,  solid  residues.  This  shows  how  remark- 
ably clean  the  solution  remained.  The  0.14  per  cent,  of  sul- 


LIXIVIATION  WITH  SODIUM  HYPOSULPHITE  177 

phuric  acid  corresponds  exactly  with  a  saturated  solution  of 
gypsum  (1  part  in  400  parts  of  water).  The  0.098  per  cent, 
chlorine  is  equal  to  0.16  per  cent,  sodium  chloride. 

Best  Strength  of  Solution.  —  In  my  experience  I  have  found 
the  best  strength  of  the  solution  to  be  0.25  to  0.50  per  cent. 
Such  a  solution  offers  not  only  the  above-mentioned  economical 
advantage,  but  it  produces  also  a  precipitate  much  cleaner  and 
richer  in  silver.  Lead  sulphate  and  cuprous  chloride  dissolve 
much  more  easily  in  a  strong  than  in  a  weak  solution,  and  there- 
fore with  a  strong  solution  a  very  low-grade  precipitate  will  result 
and  a  much  larger  amount  of  the  precipitant  will  be  required. 
It  is  a  wrong  supposition  that  a  strong  solution  shortens  the 
leaching  time  and  extracts  the  silver  better.  The  strong  solu- 
tion, dissolving  readily  lead  sulphate  and  cuprous  chloride,  be- 
comes heavily  charged  with  these  salts  and  thereby  loses  much 
of  its  dissolving  power  for  silver  chloride,  and  therefore  neither 
the  leaching  time  will  be  shortened  nor  will  the  extraction  be 
better. 

Method  of  Leaching. — When  base-metal  leaching  is  finished  and 
the  sodium  hyposulphite  solution  is  to  be  applied,  it  is  well  to  do 
this  before  all  the  water  has  disappeared  from  the  surface  of  the 
ore.  The  solution,  th.en,  in  its  course  downward,  follows  closely  the 
water  and  replaces  the  latter  perfectly  in  all  parts  of  the  charge, 
and  has  less  opportunity  to  become  mixed  with  the  water  absorbed 
by  the  ore  than  if  the  water  is  allowed  to  drain  from  the  ore  be- 
fore the  solution  is  applied,  and  consequently  the  reaction,  when 
the  silver  appears  at  the  outlet,  will  be  more  precise.  There  is 
another  advantage  connected  with  it,  inasmuch  that  if  the  charge 
be  allowed  to  drain,  the  space  before  occupied  by  water  will  be 
filled  by  air,  and  when  the  hyposulphite  solution  is  run  on,  it  is 
very  difficult,  in  fact  almost  impossible,  to  drive  this  air  out  again. 
The  layer  next  to  the  surface  will  give  up  its  air,  which  can  be 
observed  by  the  bubbles  at  the  beginning,  but  further  down  in 
the  charge  the  friction  becomes  too  great,  and  the  air,  not  being 
able  to  escape,  will  compress  to  let  the  solution  pass,  and  by 
doing  so  will  reduce  the  speed  of  filtration  and  prevent  a  quick 
and  intimate  contact  between  the  ore  particles  and  solution, 
thus  increasing  the  time  of  leaching  and  decreasing  the  percentage 
of  extraction.  Therefore  attention  should  be  paid  that  during 
no  time  of  lixiviation  the  liquid  be  allowed  to  sink  below  the 


178  HYDROMETALLURGY  OF  SILVER 

surface  of  the  ore.  In  a  fresh  charge  the  ore  is  loose,  and  when 
water  is  applied  on  top  it  will,  in  descending,  easily  force  down- 
ward the  air,  which  will  escape  through  the  outlet. 

Testing  the  Solution  for  Silver.  —  It  requires  close  watching 
to  ascertain  the  time  when  the  silver  appears  at  the  outlet.  If 
the  liquid  is  tested  with  sodium  or  calcium  sulphide  the  reaction 
is  very  uncertain,  because  the  base-metal  leaching  is  seldom 
carried  to  such  an  extent  that  no  light  clouds  are  formed  by  an 
addition  of  these  reagents,  and  therefore  the  reaction  cannot 
show  the  first  traces  of  silver  in  the  outflowing  stream.  Even  a 
very  dilute  solution  of  sodium  hyposulphite  dissolves  silver 
chloride,  and  we  can  safely  assume  that,  as  soon  as  sodium  hypo- 
sulphite can  be  detected  at  the  outlet,  the  stream  contains  silver; 
and  since  furthermore  the  liquid  naturally  will  contain  more 
hyposulphite  salt  than  silver,  it  is  safer  to  adopt  a  method  by 
which  this  salt  can  be  detected  in  the  outflowing  stream.  I  use 
the  following  test,  which  is  reliable  and  convenient: 

A  small  strip  of  starch  paper  is  dipped  into  iodine  solution 
and  then  held  in  the  stream.  If  the  blue  color  disappears  it  is 
a  sign  that  the  liquid  contains  a  hyposulphite  salt  and  conse- 
quently silver,  and  the  stream  has  to  be  turned  at  once  into  the 
trough  leading  to  the  precipitation  tanks.  The  base  metals  have 
to  be  leached  with  cold  water  to  make  this  test  applicable,  because 
hot  water  also  discolors  the  blue  paper.  I  advise  all  who  prac- 
tise the  lixiviation  process  to  use  this  test  and  see  that  when 
base-metal  leaching  is  changed  to  silver  leaching  the  outflowing 
stream  is  very  closely  watched.  By  doing  so  they  will  find 
that  shortages  in  silver  for  which  they  could  not  account  before 
will  be  avoided. 

Effect  of  Lead  and  Copper.  —  If  the  ore  contains  lead  and 
copper  we  shall  find  both  metals  in  the  hyposulphite  solution, 
because  lead  sulphate  and  cuprous  chloride  which  are  present 
in  the  roasted  ore  are  not  soluble  in  water  but  dissolve  in  sodium 
hyposulphite.  Both  these  metals  are  precipitated  together  with 
the  silver,  and  we  find  them  in  large  quantities  in  the  precipitate, 
reducing  materially  the  fineness  of  the  latter.  To  remove  the 
lead  from  the  solution  Mr.  Russell  precipitates  it  as  carbonate  by 
adding  sodium  carbonate  to  the  solution  previous  to  precipitation 
with  sodium  sulphide.  By  doing  so  he  obtains  a  sulphide  pre- 
cipitate free  from  lead  and  lead  carbonate  as  a  by-product,  but 


LIXIVIATION  WITH  SODIUM  HYPOSULPHITE  179 

this  complicates  the  manipulations  without  offering  much  prac- 
tical advantage.  Only  where  the  sulphides  are  refined  by  melt- 
ing them  with  iron  and  borax  in  crucibles  is  the  use  of  this  method 
justified,  because  if  the  precipitate  is  free  from  lead  a  silver 
bullion  over  900  fine  will  result.  But  this  treatment  is  too  ex- 
pensive and  inconvenient,  and  is  not  used  except  in  small  works, 
and  then  only  in  exceptional  cases.  In  large  works  the  refining 
is  done  on  a  lead  bath  in  the  cupeling  furnace,  and  lead  in  the 
precipitate  is  then  of  great  advantage. 

For  the  sake  of  information  the  Russell  method  was  tried  in  a 
modified  way.  Instead  of  precipitating  the  lead  in  the  solution 
after  it  had  left  the  lixiviation  vat,  it  was  precipitated  inside  the 
ore  charge  by  adding  sodium  carbonate  to  the  solution  before 
leaching.  The  outflowing  solution  was  entirely  free  from  lead. 
The  precipitated  lead  carbonate  remained  in  the  ore.  No  in- 
jurious effect  on  the  extraction  was  noticed.  The  residues  con- 
tained about  as  much  silver  as  when  leached  without  an  addition 
of  sodium  carbonate,  but  the  precipitate  was  much  richer  in 
silver. 

Calcium  Sulphide  as  Precipitant,  and  Action  of  Calcium  Hypo- 
sulphite. —  If  calcium  sulphide  is  used  as  precipitant  the  solution 
will  contain  after  precipitation  calcium  hyposulphite,  and  it 
used  to  be  generally  assumed  that  within  a  short  time  the  original 
sodium  hyposulphite  solution  was  replaced  by  calcium  hyposul- 
phite. I  discovered  and  demonstrated  that  this  is  not  the  case; 
in  fact,  that  even  if  the  original  lixiviating  solution  were  calcium 
hyposulphite,  that  compound  could  not  exist  in  the  process  for 
any  length  of  time,  but  would  be  converted  into  sodium  hyposul- 
phite. We  read  in  standard  works  that  Patera  leached  with 
sodium  hyposulphite  and  precipitated  with  sodium  sulphide, 
while  Kiss  used  calcium  hyposulphite  as  solvent  and  calcium 
sulphide  as  precipitant.  It  is  mentioned  that  calcium  hyposul- 
phite possesses  a  greater  dissolving  energy  for  gold  than  the  cor- 
responding sodium  salt,  and  that  for  this  reason  the  Kiss  process 
is  more  suitable  for  gold-bearing  silver  ores.  However,  Kiss  was 
not  leaching,  as  he  thought,  with  calcium  hyposulphite,  but  with 
sodium  hyposulphite,  and  the  better  extraction  of  gold  was 
most  likely  due  to  cooling  the  roasted  ore  slowly. 

In  roasting  sulphureted  ore  with  salt,  sodium  sulphate  is 
formed  in  large  quantities.  This  salt  is  not  easily  removed  from 


180  HYDROMETALLURGY  OF  SILVER 

the  ore  by  leaching  with  water.  If  leaching  with  water  be  pro- 
longed it  will  be  found  that  the  outflowing  liquid  will  react  for 
sodium  sulphate  long  after  all  the  heavy  metallic  salts  are  re- 
moved. Therefore,  when  the  hyposulphite  solution  for  the  ex- 
traction of  silver  is  applied  the  ore  still  contains  sodium  sulphate 
in  considerable  quantity.  Calcium  hyposulphite  reacts  with 
sodium  sulphate,  forming  sodium  hyposulphite  and  insoluble 
calcium  sulphate.  If,  therefore,  calcium  hyposulphite  is  brought 
into  the  stock  solution  by  the  precipitant,  or  when  calcium  hypo- 
sulphite is  used  as  solvent,  it  will  be  converted  into  sodium  hypo- 
sulphite by  the  regular  operation  of  the  process. 

Time  Required  for  Lixiviation.  —  The  time  of  lixiviation  varies 
according  to  the  nature  of  the  ore,  its  permeability,  and  the  size 
of  the  charge.  I  observed  that  if  the  main  part  of  the  silver  in 
the  ore  is  contained  in  galena,  the  silver  extraction  will  be  slow, 
even  if  the  ore  filters  well,  while  if  all  or  the  principal  part  of  the 
silver  is  contained  in  copper,  zinc,  arsenical  or  antimonial  minerals 
the  extraction  is  quick.  At  the  Silver  King  mine,  Arizona,  a 
charge  of  8  tons  required  only  nine  hours'  silver  leaching,  though 
it  contained  on  an  average  161.4  oz.  silver  per  ton;  while  at  Cusi- 
huiriachic,  Mexico,  the  time  required  for  a  charge  of  8  tons  was 
fifty-three  hours,  the  ore  containing  but  47  oz.  silver  per  ton. 
The  filtering  capacity  in  both  cases  was  very  nearly  the  same. 
At  the  Silver  King  the  silver  was  mostly  contained  in  gray  copper 
ore,  antimonial  fahlerz  and  silver-copper  glance,  while  at  Cusi- 
huiriachic  it  was  principally  contained  in  galena.  I  have  made 
the  same  observation  with  ores  of  many  other  localities. 

A  free  filtration  is  very  important  for  a  quick  and  thorough 
extraction.  The  ore  particles  should  be  brought  rapidly  in  con- 
tact with  fresh  solution,  which  cannot  be  done  if  the  filtration  is 
slow.  After  the  solution,  descending  through  the  ore,  has  dis- 
solved a  certain  quantity  of  the  salts  present  it  loses  much  of  its 
dissolving  energy,  and  therefore  the  small  stream  of  a  bad  filter- 
ing ore  will  not  contain  much  more  silver  per  liter  than  the  large 
stream  of  a  quick-filtering  ore.  The  same  observation  can  be 
made  if  the  outflowing  stream  of  a  quick-filtering  ore  be  checked 
and  reduced  to  a  small  stream.  Likewise,  if  lixiviation  is  in- 
terrupted and  the  solution  allowed  to  remain  in  contact  with  the 
ore  for  some  time,  say  over-night,  the  solution  will  not  contain 
much  more  silver  after  leaching  is  resumed  than  the  stream  did 


LIXIVIATION  WITH  SODIUM  HYPOSULPHITE  181 

before  the  interruption.  Ores  which  after  roasting  run  on  the 
cooling  floor  like  water  always  filter  badly  and  are  not  suitable 
for  tank  lixiviation.  Mixing  such  ores  with  sand  or  chopped 
straw  does  not  improve  their  permeability.  By  mixing  in  a 
small  percentage  of  galena  before  roasting,  however,  the  ore 
loses  somewhat  of  its  dusty  condition  and  permits  a  little  better 
percolation.  The  only  rational  way,  however,  of  treating  such 
ores  is  by  trough  lixiviation. 

Regeneration  of  a  Foul  Hypo  Solution.  —  The  same  sodium 
hyposulphite  solution  was  used  at  the  Silver  King  works,  Arizona, 
for  over  a  year  and  a  half  without  requiring  any  addition  what- 
soever, of  sodium  hyposulphite,  and  acted  still  as  energetically  as 
it  did  the  first  day.  The  ore  permitted  a  very  quick  extraction 
of  the  silver;  the  silver  leaching  did  not  require  more  than  nine 
to  ten  hours.  Suddenly  a  change  took  place.  The  solution, 
which  formerly  gained  in  strength  and  had  to  be  diluted  from 
time  to  time,  became  weaker;  the  time  required  for  silver  leaching 
increased  to  fifty,  sixty,  and  seventy  hours.  The  main  portion 
of  the  silver  was  extracted  in  twenty-four  hours,  but  the  balance 
took  a  very  long  time.  The  value  of  the  precipitate  dropped  from 
$12  per  pound  to  $6,  and  the  leaching  tanks,  which  always  were 
ahead  of  the  furnaces,  could  not  keep  up  with  them.  An  addition 
of  sodium  hyposulphite  salt  to  the  stock  solution  improved  it 
somewhat,  but  only  for  a  short  time.  Seeking  for  the  cause  of 
this  strange  change,  I  examined  closely  the  present  ore  and 
compared  it  with  that  which  was  worked  before,  as  the  cause 
could  be  found  only  there,  because  the  roasting  as  well  as  the 
leaching  was  done  in  exactly  the  same  way  as  during  the  previous 
year  and  a  half.  I  found  that  the  present  ore,  though  its  general 
appearance  was  the  same  as  before,  did  not  carry  any  copper. 
This  occurred  when  the  works  at  the  mine  reached  the  700-ft. 
level.  Not  being  able  to  find  any  other  difference,  I  suspected  the 
absence  of  copper  in  the  ore  to  be  the  cause  of  the  fouling  of 
the  solution,  and  after  an  experiment  on  a  small  scale  had  proved 
the  supposition  to  be  correct,  cupric  chloride  was  prepared  by 
boiling  blue  vitriol  with  salt.  One  precipitation  tank  was  cleaned 
and  filled  with  stock  solution.  Then  12  gallons  of  the  prepared 
cupric  chloride  solution  was  added,  stirred  and  precipitated  with 
calcium  sulphide.  After  precipitation  the  solution  of  this  tank 
was  capable  of  dissolving  nearly  50  per  cent,  more  silver  chloride 


182  HYDROMETALLURGY   OF  SILVER 

than  before.  This  was  tried  in  the  laboratory  with  freshly  pre- 
pared silver  chloride.  The  clear  solution  was  decanted  from  the 
copper  precipitate  and  conveyed  to  the  pump  tank  to  be  mixed 
with  the  balance  of  the  stock  solution.  During  the  first  two 
days  the  copper  precipitations  were  made  daily.  A  change  in 
the  action  of  the  solution  could  be  observed  at  once.  In  all 
nine  copper  precipitations  were  made,  with  a  consumption  of 
127  Ib.  of  bluestone  and  181  Ib.  of  salt.  At  the  end  of  the  ninth 
precipitation  the  stock  solution  had  almost  regained  its  former 
dissolving  energy.  The  silver  leaching  time  dropped  to  fourteen 
and  sixteen  hours,  and  the  value  of  the  precipitate  improved 
again  to  $10  and  $11  per  pound.  The  solution  kept  in  this 
excellent  condition  for  over  a  month,  when  signs  of  degeneration 
could  be  observed;  a  fresh  precipitation,  however,  brought  it 
right  again.  While  not  being  prepared  to  offer  a  chemical  expla- 
nation why  the  want  of  copper  in  the  ore  should  cause  a  degen- 
eration of  the  sodium  hyposulphite  solution,  and  why  an  addition 
of  cupric  chloride  to  this  solution,  which  is  precipitated  before  it  is 
sent  in  circulation,  should  restore  to  the  solution  its  former  quality 
and  dissolving  energy,  I  place  my  observation  on  record  with  the 
belief  that  it  will  be  of  great  service  to  many  lixiviation  works. 

In  Cusihuiriachic,  Mexico,  it  took  seventy-two  hours  to  leach 
the  silver  from  a  tank  charge  of  8  tons,  while  after  a  treatment 
of  the  stock  solution  with  cupric  chloride  it  was  accomplished  in 
twelve  to  eighteen  hours. 

Filters  for  Leaching  Tanks.  —  A  free  filtration  being  of  great 
importance  for  the  success  of  the  process,  the  selection  of  a  proper 
filter  is,  therefore,  also  of  great  importance.  Some  use  gravel 
and  sand  without  a  wooden  filter  bottom,  as  at  La  Baranca, 
Sonora,  Mexico.  The  material  of  such  a  filter  is  cheap  enough, 
but  its  preparation  is  troublesome,  and  therefore  a  filter  is  usually 
kept  in  use  until  the  filtration  becomes  so  bad  that  it  has  to  be 
replaced  by  a  new  one.  At  Sombrerete  straw  filters  were  once 
used.  Short  straw  was  spread  on  a  wooden  filter  bottom  about 
a  foot  thick  and  then  the  ore  was  dumped  on  top  of  it.  This 
rather  coarse  filter  did  not  produce  a  very  clear  filtrate  when 
new,  but  after  the  first  charge  it  improved  in  this  respect  and 
produced  a  clear  solution;  however,  the  outflowing  stream  de- 
creased with  every  charge,  and  when  the  filter  had  to  be  renewed 
it  was  found  that  the  straw  had  rotted  and  packed  tightly. 


LIXIVIATION  WITH  SODIUM  HYPOSULPHITE  183 

Filters  like  those  described  above,  which  cannot  be  cleaned, 
and  are  therefore  kept  in  use  as  long  as  possible,  invariably  cause 
a  decrease  of  the  working  capacity  of  the  leaching  vats  and  with 
it  a  decrease  in  production  and  percentage  of  extraction.  Burlap, 
the  material  used  for  making  grain  sacks,  makes  the  best  and 
most  lasting  filter.  It  is  cut  in  pieces  and  spread  over  the  wooden 
filter  bottom  so  that  one  strip  laps  over  the  other  about  3  in. 
The  ends  are  rolled  up  and  packed  tightly  around  the  circum- 
ference, or  better  into  a  groove  as  shown  at  N,  Fig.  31.  After 
each  charge  the  strips  of  burlap  are  carefully  rolled  up,  so  that 
no  residue  which  adheres  to  them  may  drop  below  the  filter 
bottom,  are  well  washed  and  again  spread  on  the  filter  bottom. 
If  this  is  done  regularly  the  filter  is  kept  in  the  best  possible  con-, 
dition.  If  the  washing  is  not  done  after  every  charge  a  thin 
layer  of  residue  will  form  on  the  filter  cloth,  gradually  growing 
thicker  and  harder,  partly  by  the  repeated  pressure  of  the  feet  of 
the  laborers  and  partly  by  the  deposition  of  calcium  sulphate. 
The  gypsum  also  incrusts  the  fibers  of  the  cloth,  makes  it  stiff 
and  hard,  and  finally  stops  filtration  entirely.  Where  burlap 
cannot  be  obtained  coarse  sheeting  is  a  good  substitute. 

End  of  the  Silver  Leaching.  —  The  extraction  of  the  silver  from 
a  charge  may  be  finished  while  the  outflowing  solution  still  gives 
a  considerable  precipitate  upon  addition  of  calcium  or  sodium 
sulphide.  It  is  impossible  to  judge  by  the  color  of  this  precipitate 
whether  it  contains  silver  or  not.  It  is,  however,  important  to 
know  when  all  soluble  silver  is  extracted,  in  order  to  avoid  unneces- 
sary consumption  of  the  precipitant  and  loss  of  time.  To  ascer- 
tain when  the  end  of  the  silver  leaching  has  been  reached  a  large 
beaker  is  filled  with  the  solution  and  some  calcium  sulphide  is 
added.  The  precipitate  is  allowed  to  settle,  the  clear  solution 
decanted,  and  the  precipitate  poured  on  a  paper  filter  and  washed 
well.  It  is  then  removed  from  the  filter  and  dissolved  in  nitric 
acid,  filtered  to  remove  the  sulphur,  and  a  drop  or  two  of  hydro- 
chloric acid  is  added.  If  a  white  precipitate  or  only  a  cloudiness 
is  produced,  which  by  dilution  and  boiling  does  not  disappear, 
there  is  still  silver  in  the  solution  and  the  leaching  has  to  be 
continued.  If  the  solution  remains  clear  the  extraction  is  con- 
cluded. 

The  influx  of  sodium  hyposulphite  is  then  stopped  and  the 
solution  is  allowed  to  drain  until  it  commences  to  disappear 


184  HYDROMETALLURGY  OF  SILVER 

under  the  surface  of  the  ore,  when  a  stream  of  water  is  turned  on 
to  displace  the  sodium  hyposulphite  solution  absorbed  by  the 
ore.  It  is  only  necessary  to  continue  this  second  application  of 
water  for  a  comparatively  short  time;  just  long  enough  to  keep 
the  same  volume  of  stock  solution  on  hand.  Then  the  charge  is 
allowed  to  drain  as  dry  as  circumstances  permit,  after  which  the 
residues  are  discharged.  This  is  done  by  shoveling  them  into 
chutes,  of  which  one  is  placed  between  each  two  tanks  and  which 
discharge  into  cars  running  underneath  the  tanks  (Fig.  31),  or 
into  a  large  triangular  trough  beneath  the  vats,  in  which  a  current 
of  water  flows.  The  trough  ought  not  to  have  an  inclination  less 
than  1  in.  to  the  foot.  In  some  works  the  residues  are  sluiced 
out.  To  make  this  method  successful  the  vats  must  not  be  too 
deep  nor  of  too  large  diameter.  The  stream  of  water  has  to  be 
applied  under  pressure  by  means  of  a  pump,  or  under  good  head 
from  a  storage  tank. 


XIV 

PRECIPITATION  OF  SILVER 

IN  all  modern  lixiviation  works  the  precipitation  vats  are 
provided  with  a  mechanical  contrivance  to  agitate  the  solution 
during  and  after  precipitation,  and  only  in  antiquated  works  is 
this  operation  done  by  hand  with  a  paddle.  A  horizontal  beam 
about  12  in.  shorter  than  the  inside  diameter  of  the  vat  is  fastened 
to  a  vertical  iron  shaft  (Figs.  57  and  58).  This  horizontal  beam, 
which  moves  above  the  surface  of  the  solution,  is  provided  with 
hard-wood  staves  about  2  in.  square  and  reaching  down  to  about 
1J  in.  above  the  bottom.  These  staves  are  so  arranged  that 
when  the  agitator  is  in  motion  they  cut  the  liquid  with  the  edge. 
The  agitator  is  set  in  motion  by  a  friction  clutch,  and  it  should  be 
started  gradually  to  avoid  breaking  the  staves.  The  inside  of 
the  tank  is  provided  with  four  vertical  wooden  wings,  projecting 
3  in.  toward  the  center  and  reaching  nearly  to  the  bottom.  They 
break  the  violent  current  around  the  periphery  and  throw  the 
solution  toward  the  center,  thus  causing  a  strong  whirling  motion. 
The  agitator  should  make  30  r.p.m.  if  the  diameter  of  the  vat 
is  not  more  than  8  or  9  ft.  Agitators  of  this  construction  are 
only  durable  when  used  in  vats  not  deeper  than  6  ft.  In  deeper 
tanks  the  staves  will  break,  and  an  agitator  of  stronger  construction 
has  to  be  used. 

Another  method  of  agitation  which  is  very  convenient  is  by 
the  use  of  compressed  air,  furnished  by  a  small  compressor.  From 
the  receiver  a  pipe-line  leads  along  the  whole  row  of  precipitation 
vats.  At  each  vat  is  inserted  a  T  with  a  f-in.  branch.  These 
branch  tubes  are  provided  with  valves  and  connected  with  a 
rubber  hose  6  to  8  ft.  long,  the  other  end  of  which  terminates 
with  a  f-in.  gas  pipe.  This  pipe  must  be  long  enough  to  reach 
any  point  of  the  bottom  and  still  project  about  2  ft.  above  the 
rim  of  the  tank.  If  the  vat  is  full  and  the  air  is  turned  on  the 
solution  is  put  into  violent  motion,  as  if  boiling.  It  is  well  to 

185 


186  HYDROMETALLURGY   OF  SILVER 

change  the  position  of  the  pipe  from  time  to  time,  especially  if 
the  vat  has  a  large  diameter.  This  method  of  agitating  the 
solution  is  now  extensively  used. 

Preparing  Calcium  Poly  sulphide.  —  This  precipitant  is  pre- 
pared by  boiling  milk  of  lime  with  pulverized  sulphur.  The 
proportion  has  to  be  ascertained  by  the  operator,  as  it  depends 
on  the  quality  of  the  burnt  lime  rock.  No  free  sulphur  should 
be  visible  in  the  white  residues.  It  is  soon  found  out  how  much 
of  the  local  lime  has  to  be  slacked  for  a  certain  weight  of  sulphur. 

The  boiling  was  formerly  done  in  upright  boilers,  made  of 
boiler  iron  4  ft.  in  diameter  and  6  to  7  ft.  deep,  by  direct  applica- 
tion of  a  steam  jet.  Near  the  bottom  was  a  pipe  outlet  closed 
with  a  wooden  plug.  This  outlet  discharged  into  an  iron 
tank  in  which  a  sand  filter  was  prepared.  Below  the  level  of  the 
bottom  of  this  filter  tank  was  placed  the  iron  storage  tank  to 
receive  the  calcium  sulphide  solution.  From  this  storage  tank  a 
pipe-line  conveyed  the  chemical  to  the  precipitation  tanks.  The 
boiling  tank  is  first  charged  with  the  milk  of  lime,  but  enough 
margin  should  be  left  for  the  water  condensed  from  the  steam, 
which  in  the  beginning  is  quite  considerable.  When  the  milk  is 
boiling  the  sulphur  is  charged  gradually,  by  means  of  a  sieve,  in 
order  to  scatter  the  same  as  finely  as  possible  over  the  surface. 
If  charged  by  a  shovel  the  sulphur  sinks  quickly  to  the  bottom, 
forming  chunks  which  require  a  long  time  to  combine  with  the 
lime.  When  boiling  is  concluded  some  cold  water  is  splashed 
over  the  surface,  which  makes  the  foam  disappear  and  also  causes 
a  quicker  settling.  When  settled  the  clear  yellow-brown  solution 
is  decanted,  by  means  of  a  stiff  1-in.  rubber  hose,  into  the  storage 
tank.  This  done,  the  plug  near  the  bottom  of  the  boiler  is  re- 
moved, and  the  sediment  allowed  to  flow  into  the  filter  tank. 
To  make  it  flow  more  freely,  some  water  is  added  before  removing' 
the  plug.  The  filtrate  from  the  residues  also  flows  into  the  storage 
tank.  These  residues  are  very  slow  in  filtering  and  retain  much 
calcium  sulphide  solution,  which  has  to  be  displaced  by  applying 
water  on  top  of  the  residues,  which,  however,  is  rather  difficult 
to  do  properly,  as  the  residues  are  very  pasty  and  very  slow 
filtering,  so  that,  when  they  are  removed  by  shovels,  some  of  the 
calcium  sulphide  will  be  wasted. 

The  calcium  sulphide  department  was  much  improved  lately 
by  myself.  Fig.  35  represents  a  vertical  section  of  the  arrange- 


PRECIPITATION   OF  SILVER 


187 


188 


HYDROMETALLURGY   OF  SILVER 


ment.  A  is  the  lime-slacking  box,  nearly  level  with  the  cool- 
ing floor.  When  the  lime  is  slacked  the  milk  is  allowed  to  flow 
into  the  distributing  trough  B,  whence  it  fills  the  boiler  C. 
There  are  two  of  these  boilers,  side  by  side,  which  can  be  alter- 
nately charged  through  the  distributing  trough.  This  trough  is 
made  of  J-in.  steel  and  is  represented  by  Fig.  36.  The  outlets 
in  the  bottom  can  be  closed  by  wooden  plugs. 

The  boiler  C  is  represented  in  detail  by  Fig.  37.  The  sides 
are  made  of  f-in.  while  the  bottom  and  top  are  of  f-in.  steel. 
It  is  a  boiler  and  pressure  tank  combined.  Fig.  38  gives  a  top 
view.  There  are  six  openings  in  the  top,  of  which  one  is  a  man- 
hole, while  the  other  five  are  smaller,  and  are  connected  with 


~~l    t 

- 

--0—0- 

"* 

FIG.  36.  — DISTRIBUTING  TROUGH  FOR  MILK  OF  LIME. 
To  be  made  of  steel. 

pipes.  Three  of  them  reach  nearly  to  the  bottom  (Fig.  35),  viz.: 
the  discharge  pipe  passing  through  the  center,  the  steampipe, 
and  the  compressed-air  pipe.  The  filling  and  air-escape  pipes  do 
not  extend  into  the  boiler.  The  reason  why  the  compressed-air 
pipe  reaches  nearly  to  the  bottom  is  to  agitate  the  pulp  during 
discharging.  The  manhole  cover  has  a  4-in.  flanged  opening  in 
the  center  (Fig.  39),  which  can  easily  be  closed  by  a  cast-iron 
plate  and  bolts.  This  opening  serves  for  introducing  the  sulphur; 
it  is  left  open  during  filling  and  boiling,  and  is  closed  during 
discharging. 

When  the  milk  of  lime  is  charged  steam  is  turned  on.     The 
opening  in  the  manhole  cover  is  kept  open  for  the  escape  of  the 


PRECIPITATION   OF  SILVER 


189 


'16 


FIG.  37. —BOILER  AND  PRESSURE  TANK  FOR 

CALCIUM  SULPHIDE. 

Vertical  section. 


190 


HYDROMETALLURGY   OF  SILVER 


air  during  filling  and  for  the  escape  of  vapor  during  boiling. 
Room  has  to  be  left  for  the  water  from  condensed  steam,  and  also 
for  foaming,  which  sometimes  occurs  when  too  much  steam 


FIG.  38.— BOILER  AND  PRESSURE  TANK  FOR 

CALCIUM  SULPHIDE. 

Plan  of  head. 

is  admitted.  By  means  of  a  wooden  staff  in  which  notches  are 
cut,  and  which  is  inserted  through  the  opening  in  the  manhole 
cover,  the  filling  is  regulated. 

It  is  not  advisable,  for  two  reasons,  to  make  too  strong  a 


FIG.  39.  — BOILER  AND  PRESSURE  TANK 

FOR  CALCIUM  SULPHIDE. 
Saddle  to  be  riveted  to  head  of  tank. 

solution.  In  a  very  strong  solution  crystals  of  bisulphide  of  cal- 
cium are  formed,  which  will  be  found  on  the  sides  and  bottom  of 
the  storage  tank,  and  will  also  be  formed  in  the  pipe-line  leading 


PRECIPITATION   OF  SILVER  191 

from  that  tank  to  the  precipitation  tank,  and  before  long  will 
clog  the  pipe,  necessitating  the  taking  down  of  the  whole  line 
for  cleaning.  The  second  inconvenience  caused  by  too  strong 
a  solution  is  the  difficulty  to  see  the  end  reaction  in  precipitating, 
in  consequence  of  which  the  precipitator  is  apt  to  add  too  much 
of  the  precipitant. 

I  found  that  a  solution  answers  well  if,  in  boiling,  to 
each  cubic  foot  of  water  about  2J  Ib.  of  sulphur  are  taken. 
Of  good  and  freshly  burned  lime  about  2i  Ib.  to  the  pound 
of  sulphur  is  an  average  proportion,  but,  as  stated  above,  it 
depends  entirely  on  the  quality  of  the  lime. 

If  it  is  intended  to  fill  the  boiler  three-quarters  full,  the  cubic 
content  is  calculated  and  by  it  the  amount  of  required  sulphur 
and  lime  ascertained.  The  lime  is  then  slacked  and  the  milk 
charged  into  the  boiler,  then  water  is  added  to  fill  the  boiler 
three-quarters  full.  Then  steam  is  admitted.  The  sulphur  is 
charged  gradually  when  the  water  commences  to  boil.  Boiling 
has  to  be  continued  for  four  or  five  hours,  according  to  circum- 
stances. If  in  a  sample,  taken  with  a  long-handled  iron  ladle  from 
the  bottom  after  filtering,  some  free  sulphur  should  be  found, 
boiling  should  be  continued.  If  this  has  no  effect  it  shows  that 
lime  is  wanting,  which  has  to  be  added.  In  the  next  charge 
more  lime  has  to  be  taken  in  the  beginning. 

When  boiling  is  concluded  the  steam  is  turned  off  and  the 
volume  of  the  solution  measured  with  the  wooden  staff;  if  the 
volume  is  short,  it  has  to  be  filled  up  with  water  to  the  mark  in 
the  staff.  Then  the  opening  in  the  manhole  cover  is  closed  tight 
and  compressed  air  is  applied.  The  discharge  pipe  is  connected 
with  a  24-in.  filter  press,  as  shown  in  Fig.  35.  The  filtrate  flows 
into  the  calcium  sulphide  storage  tank,  and  from  there  is  conveyed 
along  and  above  the  precipitation  tanks,  which  are  also  shown  in 
this  figure.  When  the  filter  press  is  filled  with  residues,  water 
under  pressure  is  forced  into  the  press  to  wash  the  residues. 
The  wash-water  is  allowed  to  run  into  the  storage  tank  and  mix 
with  the  strong  solution.  In  the  floor  near  the  lime-box  is  a 
chute,  discharging  into  iron  cars.  The  residues  from  the  lime- 
box  as  well  as  from  the  filter  press  are  thrown  into  this  chute. 

Adding  the  Precipitant.  —  The  precipitant,  whether  sodium  or 
calcium  sulphide,  is  kept  in  a  reservoir  made  of  boiler  iron,  from 
which  it  is  conveyed  through  an  iron  pipe  to  the  precipitation 


192  HYDROMETALLURGY  OF  SILVER 

vats.  At  each  vat  there  is  attached  to  the  pipe  a  hose,  which 
is  closed  by  a  pinch-cock.  In  commencing  to  precipitate  it  is 
well  to  open  the  hose  a  little  and  to  throw,  by  swinging  it,  some 
of  the  precipitant  over  the  surface.  By  the  appearance  of  the 
clouds  which  are  formed,  whether  heavy  or  light,  the  precipitator 
can  see  at  once  if  he  has  to  precipitate  a  concentrated  or  dilute 
charge.  Then  he  sets  the  solution  in  agitation  and  allows  the 
precipitant  to  flow  in.  An  experienced  precipitator  can  judge 
by  the  color  which  is  created  when  he  splashes  some  of  the  pre- 
cipitant over  the  surface,  while  the  solution  is  agitated,  the 
progress  of  precipitation,  and  knows  when  it  is  nearing  the  end. 

While  precipitation  is  going  on  the  clouds  which  are  formed 
become  gradually  lighter  in  color,  and  toward  the  end  almost 
yellow.  When  nearly  finished  the  influx  of  the  precipitant  is 
stopped,  and  after  a  few  minutes  the  agitation  also.  Then  the 
flaky  precipitate  is  allowed  to  sink  somewhat  below  the  surface, 
and  some  of  the  precipitant  is  splashed  over  the  surface.  Accord- 
ing to  the  appearance  of  the  clouds  more  or  less  precipitant  is 
added  and  the  solution  is  agitated  again.  This  operation  is  re- 
peated until  precipitation  is  complete.  If  by  an  addition  of  the 
precipitant  no  reaction  takes  place,  it  is  well  to  throw  some  strong 
silver  solution  over  the  surface  after  the  precipitate  has  partially 
settled.  If  the  places  where  the  silver  solution  fell  turn  reddish 
brown,  the  precipitant  is  in  excess  and. more  silver  solution  has 
to  be  added.  One  who  is  not  experienced  would  best  make  this 
test  in  a  beaker. 

The  precipitation  tank  as  illustrated  in  Fig.  35  is  12  ft.  in 
diameter  and  8  ft.  deep.  The  solution  is  agitated  by  compressed 
air.  The  two  outlet  pipes,  P  and  N,  are  of  lead.  To  the  upper 
pipe,  P,  is  attached  on  the  inside  of  the  tank  the  decanting  hose, 
M,  with  the  float,  S.  This  should  be  a  very  stiff  hose.  It  is  kept 
above  the  solution,  while  the  tank  is  filling  or  precipitation  is 
going  on,  by  a  thin  rope  fastened  to  a  hook  on  the  outside  of  the 
tank.  When  the  precipitate  has  settled,  the  rope  is  unhooked 
and  the  end  of  the  hose  allowed  to  float.  The  float  is  so  arranged 
that  it  keeps  the  end  of  the  hose  immersed.  When  the  float  comes 
in  near  approach  to  the  precipitate  close  attention  has  to  be 
paid,  so  that  no  precipitate  is  carried  out.  As  the  decanted  solu- 
tion is  to  be  used  again  for  extracting  silver,  it  is  conveyed  by 
means  of  troughs,  running  along  in  front  of  the  precipitation 


PRECIPITATION   OF  SILVER  193 

tanks  or  through  a  pipe-line  to  the  lower  storage  tanks,  in  which 
it  is  collected  and  elevated  to  the  upper  storage  tanks,  which  are 
placed  on  a  higher  level  than  the  rim  of  the  leaching  tanks.  The 
outer  end  of  the  pipe  P  is  connected  with  a  short  piece  of  hose, 
which  lies  in  the  trough,  if  troughs  are  used,  to  avoid  splashing. 
The  lower  outlet  pipe,  N,  is  provided  with  a  valve  and  con- 
nected with  an  iron  pipe-line,  common  to  all  precipitation  tanks, 
which  leads  to  a  pressure  tank,  by  means  of  which  the  solution 
is  forced  into  a  filter  press.  The  upper  hose  P1  serves  for  ad- 
mitting the  precipitant. 


XV 

TREATMENT  OF  THE  PRECIPITATE 

The  Precipitate.  —  In  precipitating  the  base-metal  solution 
we  have  seen  that  not  all  the  metals  present  are  equally  affected 
by  the  sulphur  of  the  precipitant,  and  that  the  silver  especially 
is  more  readily  precipitated  than  the  other  metals.  This  is  also 
the  case  if  these  metals  are  dissolved  in  the  sodium  hyposulphite 
solution,  and  therefore  the  precipitate  which  is  obtained  in  the 
earliest  stage  of  precipitation  contains  far  more  silver  than  that 
obtained  later.  Thus  the  operator  has  it  in  his  power  to  make 
different  grades  of  precipitate.  This,  however,  does  not  offer 
such  advantages  as  in  base-metal  leaching,  and  is  of  no  direct 
practical  value,  because  in  order  to  maintain  the  dissolving  energy 
of  the  sodium  hyposulphite  solution  it  is  absolutely  necessary  to 
precipitate  as  perfectly  as  possible  all  the  metals  dissolved  in  it; 
but  it  explains  why  the  black  layer  of  sulphides,  which  we  fre- 
quently find  deposited  on  the  surface  of  the  ore  charge,  is  so 
much  poorer  in  silver  than  the  precipitate  itself.  If  the  precipi- 
tation was  done  well,  and  ample  time  was  given  to  the  precipitate 
to  settle,  and  the  decantation  of  the  solution  was  always  performed 
properly  and  without  mishap,  we  should  not  find  any  black  deposit 
on  the  top  of  the  ore  charge;  but  such  exact  work  is  not  always 
done,  especially  in  large  works,  which  seldom  have  a  sufficient 
number  of  vats  to  give  the  precipitate  ample  time  to  settle. 

The  different  sulphides  settle  according  to  their  respective 
specific  gravities.  Lead,  silver,  and  copper  go  down  first,  while 
antimony,  zinc,  iron,  and  free  sulphur  follow.  While  this  sepa- 
ration is  not  theoretically  perfect,  it  takes  place  to  such  a  degree 
that  the  particles  which  settle  last  may  contain  but  30  to  100  oz. 
silver  per  ton,  while  the  total  precipitate  may  contain  5000  to 
15,000  oz.  per  ton. 

This  black  layer  of  sulphides  deposited  on  top  of  the  ore  charge 

194 


TREATMENT  OF  THE  PRECIPITATE  195 

after  an  extended  lixiviation,  being  so  much  poorer  than  the 
precipitate,  does  not  involve  any  notable  loss  of  silver  if  it  is 
carefully  scraped  off  before  the  residues  are  discharged  and  the 
scrapings  are  mixed  with  the  ore  and  roasted. 

F.  Sustersic's  Method  of  Preparing  the  Precipitate  for  Refining 
by  Cupellation.  —  If  a  precipitate  contains  a  large  percentage  of 
copper,  the  refining  of  the  same  by  cupellation  with  lead  requires 
a  large  amount  of  lead,  causes  the  formation  of  a  large  amount 
of  rich  by-products,  and  increases  the  cost  of  refining. 

For  the  treatment  of  such  a  precipitate  F.  Sustersic  devised 
a  method  by  which  he  extracts  the  copper  first,  leaving  the 
precipitate  in  an  excellent  condition  for  cupellation.  Reporting 
on  very  interesting  experiments  made  with  such  precipitate,  he 
says: 

"The  crude  precipitate  produced  by  lixiviating  the  ores  of 
Avino,  Durango,  Mexico,  with  sodium  hyposulphite  was  analyzed 
and  shown  to  contain: 

Silver 4.206  per  cent. 

Gold  . . 0.2064  per  cent. 

Copper 14.80  per  cent. 

Lead 16.40  per  cent. 

Iron    •  .  .  .   0.70  per  cent. 

Zinc   0.50  per  cent. 

Chlorine 5.10  per  cent. 

Lime 3.87  per  cent. 

Sulphur 39.40  per  cent. 

Insoluble 4.20  per  cent. 

Not  ascertained    10.80  per  cent. 

"This  precipitate  is  rather  base,  which  was  caused  by  the 
fact  that  the  ore  is  very  susceptible  to  heat  and  readily  loses 
considerable  silver  by  volatilization  at  only  a  moderate  roasting 
temperature,  and  therefore  had  to  be  roasted  at  an  extremely 
low  heat,  at  whicji  practically  all  base-metal  salts  remained  in 
the  roasted  ore.  To  refine  such  a  precipitate  with  sulphuric  acid 
would  be  too  expensive  in  Avino,  and  the  refining  with  lead  by 
cupellation  would  offer  considerable  difficulties  on  account  of  the 
large  percentage  of  copper,  so  it  became  necessary  to  find  a 
proper  method  by  which  the  copper  could  be  removed  from  the 
precipitate  before  the  latter  is  subjected  to  cupellation. 

"  In  roasting  a  material  containing  copper  sulphide  and  silver 
sulphide  the  copper  is  converted  into  sulphate  at  a  very  low 
heat,  while  it  takes  a  bright-red  heat  to  convert  the  silver  into 


196  HYDROMETALLURGY  OF  SILVER 

sulphate.1  In  accordance  with  this  difference  in  the  property  of 
these  two  metals  I  made  several  tests  and  obtained  very  gratifying 
results. 

"  A  six-inch  roasting  dish  containing  100  grams  of  the  precipi- 
tate was  placed  into  a  muffle  the  temperature  of  which  was  kept 
below  red  heat.  As  soon  as  the  free  sulphur  commenced  to  burn 
the  roasting  dish  was  removed  from  the  muffle  and  the  charge 
continually  stirred  until  the  sulphur  flame  ceased.  Then  the 
charge  was  again  placed  into  the  muffle  and  roasted  at  so  low  a 
heat  that  the  chemically  combined  sulphur  oxidized  without 
ignition.  This  second  operation  required  only  fifteen  minutes. 
The  color  of  the  precipitate  changed  from  black  to  greenish  gray. 

Weight  of  raw  charge 100.00  grams. 

Weight  of  roasted  charge 85.40  grams. 

Loss  in  weight    14.60  per  cent. 

The  roasted  precipitate  ought  to  assay 4.925  per  cent,  silver. 

By  actual  assay  it  was  found  to  contain iJ*28()  per  cent,  silver. 

The  loss  by  volatilization,  therefore,  was Nil. 

"An  analysis  of  the  roasted  precipitate  gave  the  following 
result : 

Silver 4.9286  per  cent. 

Copper 18.50  per  cent. 

Lead 19.30  per  cent. 

Ferric  oxide    1.75  per  cent. 

Zinc   0.30  per  cent. 

Lime 8.60  per  cent. 

Sulphuric  acid    37.70  per  cent. 

Insolubles   6.80  per  cent. 

"Ten  grams  of  the  roasted  precipitate  were  leached  with 
water  until  the  filtrate  became  colorless.  No  trace  of  silver 
could  be  detected  in  the  filtrate.  The  copper  in  solution  was 
precipitated  with  zinc. 

The  10  grams  of  roasted  precipitate  contained 1.85  grams  copper. 

The  nitrate  was  found  to  contain  in  solution 1 .66  grams  copper. 

Copper  extracted  as  sulphate 89.73  per  cent. 

"  The  residues  on  the  filter  weighed  5.8  grams,  and  the  roasted 
precipitate  therefore  lost  42  per  cent,  of  its  weight  by  removing 
the  soluble  substances  with  water.  These  residues  were  analyzed 
and  found  to  contain: 

"  *  See  Chapters  on  "  Sulphating  Roasting,"  and  "Ziervogel's  process." 


TREATMENT  OF  THE   PRECIPITATE  197 

Silver 8.50  per  cent. 

Copper 1 .60  per  cent. 

Lead 33.50  per  cent. 

Iron    1.65  per  cent. 

Lime 6.00  per  cent. 

Sulphuric  acid    28.50  per  cent. 

Taking  the  loss  in  weight  into  calculation  which  the  roasted  precipitate  sus- 
tained by  leaching  with  water,  if  there  was  no  silver  dissolved  the  resi- 
dues should  contain 8.4965  per  cent,  silver. 

By  actual  assay  it  was  found  that  they  did  contain.  .  .  . 8.5000 per  cent. silver. 

Therefore,  silver  dissolved  by  leaching  with  water  was .  .  Nil. 

"These  results  show  that  by  this  method  nearly  90  per  cent, 
of  the  copper  in  the  precipitate  can  be  removed  and  recovered 
as  metallic  copper  if  the  roasting  is  properly  executed,  while  all 
the  silver  and  the  lead  remains  concentrated  in  the  residues, 
which  are  then  in  a  state  well  suited  for  cupellation. 

"  By  several  check-tests  it  was  proved  that  no  loss  of  silver 
is  sustained,  neither  in  roasting  the  precipitate  nor  in  leaching 
the  same  with  water." 

This  method  of  F.  Sustersic  is  undoubtedly  a  very  rational 
way  to  prepare  the  precipitate  for  cupellation.  Not  only  does 
it  remove  and  recover  nearly  90  per  cent,  of  objectionable  copper, 
but  also  it  removes  other  salts,  thus  reducing  the  weight  of  the 
roasted  precipitate  by  42  per  cent.,  which  is  connected  with  a 
corresponding  enrichment  in  silver  and  lead.  It  permits  the 
production  of  a  clean  litharge  and  greatly  reduces  the  formation 
of  rich  slag  or  froth.  There  is  nothing  to  diminish  the  usefulness 
of  this  method  if  the  free  sulphur  of  the  crude  precipitate  is  re- 
moved by  boiling  with  caustic  soda  instead  of  by  burning.  On  a 
commercial  scale  the  precipitation  of  the  copper  from  the  solu- 
tion is,  of  course,  done  with  scrap  iron. 

Accumulation  of  Sodium  Sulphate  in  the  Solution.  —  I  stated 
above  that  when  the  leaching  with  water  is  stopped  and  the  charge 
is  ready  for  silver  leaching  the  ore  still  contains  some  sodium 
sulphate,  which  during  silver  leaching  will  enter  the  stock  solu- 
tion. In  course  of  time  the  sodium  sulphate  will  therefore  accu- 
mulate to  such  a  degree  in  the  stock  solution  that  the  latter  will 
greatly  lose  in  dissolving  energy,  a  much  longer  leaching  time  will 
be  required,  and  the  percentage  of  extraction  will  suffer  much. 
When  such  trouble  arises  operators  usually  try  to  "freshen  up" 
the  solution  by  adding  more  sodium  hyposulphite  to  it.  This, 
however,  will  benefit  it  only  for  a  very  short  time,  and  the  same 


198  HYDROMETALLURGY  OF   SILVER 

trouble  will  appear  again.  Some  add  more  hyposulphite  every 
day,  thus  increasing  greatly  the  cost  of  extraction  without  get- 
ting the  solution  back  to  its  original  energetic  condition. 

Calcium  Sulphide  as  Precipitant.  —  The  best  method  is  to 
conduct  the  process  so  that  no  sodium  sulphate  will  accumulate 
in  the  stock  solution.  This  can  be  done  by  using  calcium 
sulphide  as  precipitant.  Sodium  sulphate  reacts  with  calcium 
sulphide,  forming  sodium  sulphide  and  calcium  sulphate,  which 
precipitates,  while  the  sodium  sulphide  acts  as  precipitant  for 
the  metal  chlorides  dissolved  in  the  solution.  On  account  of 
this  reaction,  which  is  of  so  great  an  advantage  to  the  lixiviation 
process,  I  always  advocate  the  use  of  calcium  sulphide  as  pre- 
cipitant. 

The  valuable  effect  of  this  reaction  is  especially  felt  if  ore  is 
treated  which  requires  a  large  percentage  of  salt  in  roasting, 
whereby  larger  quantities  of  sodium  sulphate  are  produced. 
Should  so  much  sodium  sulphate  enter  the  stock  solution  that  by 
the  regular  process  of  precipitation  all  the  sodium  sulphate  is  not 
decomposed,  this  salt  will  then  gradually  increase,  notwithstand- 
ing the  use  of  calcium  sulphide,  and  will  spoil  the  solution.  It  is 
therefore  advisable  to  add  at  certain  intervals  an  excess  of  cal- 
cium sulphide,  and  finish  precipitation  with  silver  solution. 
This  will  free  the  solution  entirely  from  sodium  sulphate. 

If  sodium  sulphide  is  used  this  reaction  does  not  take  place, 
and  not  only  will  all  the  sodium  sulphate  dissolved  during  silver 
leaching  accumulate  and  remain  in  the  stock  solution,  but  the 
amount  will  be  increased  from  the  precipitant,  which  always 
contains  more  or  less  sodium  sulphate.  In  such  works  it  is  well 
to  make  provision  for  the  manufacture  of  calcium  sulphide,  solely 
for  the  purpose  of  purifying  the  stock  solution  from  time  to  time. 
Treatment  of  the  Precipitate.  —  The  precipitate  should  be  dis- 
charged from  the  precipitation  tanks  every  other  day.  If  it  is 
allowed  to  accumulate  for  a  longer  time  it  changes  from  a  flaky 
condition  to  a  very  fine  powder,  which  settles  slowly  and  will 
cause  the  solution  in  circulation  to  contain  in  suspension  a  con- 
siderable amount  of  exceedingly  fine  precipitate.  The  precipi- 
tate is  discharged  into  a  tank  with  a  slowly  moving  agitator, 
varying  in  size  according  to  the  size  of  the  works,  from  which  the 
precipitate  is  drawn  into  a  pressure  tank  in  portions  as  required, 
and  thence,  by  means  of  compressed  air,  it  is  forced  into  a 


TREATMENT  OF  THE  PRECIPITATE  199 

Johnson  filter  press.  Pressure  tanks  made  of  cast-iron  will  resist 
longer  the  corroding  action  of  the  solution  than  those  made  of 
boiler  iron.  They  should  be  made  so  that  the  top  or  cover  can 
be  removed  when  required.  The  relief  pipe  should  return  to 
the  sulphide  tank  connected  with  the  agitator,  because  the  air 
when  relieved  escapes  with  great  force,  and  is  apt  to  carry  along 
some  precipitate.  Instead  of  compressed  air,  steam  may  be  used; 
but  the  pressure  will  be  limited  by  the  pressure  in  the  boiler,  and 
often  may  not  be  sufficient.  The  feed  pipe  of  the  press  should  be 
in  connection  with  water  and  air  pipes,  so  that  after  the  filter 
press  is  filled  the  precipitate  can  be  washed  by  pumping  warm 
water  through  it,  and  partially  dried  by  forcing  compressed  air 
through  it.  When  so  treated  the  precipitate  will  come  out  of 
the  press  in  hard  cakes,  permitting  a  clean  handling  of  them. 
The  press  should  be  placed  in  a  separate  room  with  cement  floor. 
Such  a  floor  is  easily  kept  clean,  and  prevents  loss  of  silver. 

Instead  of  a  pressure  tank,  a  steam  pump  can  also  be  used; 
only  attention  should  be  paid  that  the  steam  cylinder  of  the  pump 
is  larger  than  the  pump  cylinder,  so  that  a  pressure  of  150  Ib.  or 
more  can  be  produced  with  it,  if  it  should  be  necessary.  The 
pump  cylinder  should  be  provided  with  a  bronze  piston  rod  and 
liner,  because  iron  is  more  or  less  affected  by  the  solution,  and  in 
course  of  time  the  inside  of  an  iron  cylinder  becomes  very  rough. 
Much  cleaner  work  can  be  done  with  a  pressure  tank,  because 
there  is  always  more  or  less  leakage  around  the  piston  rod  of  a 
pump.  The  filtrate  is  conveyed  to  one  of  the  lower  solution 
reservoirs. 

It  was  found  more  convenient,  cleaner,  and  labor-saving  not 
to  use  an  agitating  tank  for  the  precipitate,  but  to  connect  the 
pressure  tank  direct  with  the  pipe-line  which  runs  in  front  of  the 
precipitation  tanks  and  with  which  they  are  in  communication. 
The  filling  of  the  pressure  tank  is  done  direct  from  the  precipi- 
tation tanks,  but  in  order  to  release  them  as  soon  as  possible  for 
their  regular  work,  the  pressure  tank  is  made  large  enough  to 
receive  the  whole  of  the  precipitate  of  one  tank.  It  is  made  of 
steel  and  used  in  an  upright  position. 

When  the  charge  has  been  forced  into  the  filter  press  and  the 
pressure  tank  is  empty,  the  compressed  air  contained  therein  is 
allowed  to  escape  by  opening  the  valve  of  the  air-escape  pipe. 
The  air  rushes  out  with  great  force,  carrying  out  moisture  and 


200 


HYDROMETALLURGY   OF  SILVER 


some  of  the  precipitate  hanging  on  the  side  near  the  air  outlet, 
which,  if  left  to  escape  free,  would  not  only  cause  some  loss  of 
silver,  but  make  the  surroundings  very  unclean.  To  avoid  this 
I  designed  and  introduced  a  drum  into  which  the  air  is  discharged, 
and  which  gives  very  good  satisfaction.  Fig.  40  is  a  vertical 
section.  The  drum  is  24  in.  in  diameter  and  3  ft.  6  in.  high.  In 
alternating  distances  of  6  in.,  4  in.,  6  in.,  4  in.,  etc.,  angle  irons 
A,  A,  are  riveted  to  the  inside,  forming  circular  shelves.  On 
these  shelves  rest  conical  trays,  one  with  the  cone  turned  down, 


FIG.  40.— AIR  BLOW-OFF  DRUM. 
Vertical  section. 

the  other  with  the  cone  turned  up.  Those  with  the  cone  down 
have  a  6-in.  circular  opening  in  the  center,  while  the  others  have 
openings  (C,  C,  Fig.  41)  near  the  periphery.  These  trays  are 
kept  in  place  by  four  small  bolts.  Underneath  the  bottom  tray 
is  an  outlet  pipe  leading  to  one  of  the  lower  storage  tanks,  while  at 
an  angle  of  90  deg.,  enters  the  air-escape  pipe  from  the  pressure 
tank.  The  working  of  this  air  blow-off  drum  is  clearly  explained 
by  the  drawing. 

The  Pressure  Tanks.  —  If  a  pressure  tank  is  destined  to  lift 
liquors  containing  residues  or  a  precipitate  it  ought  to  be  used 
in  an  upright  position,  because  it  facilitates  the  discharge  of  these 


TREATMENT   OF  THE   PRECIPITATE 


201 


solids,  but  if  such  a  tank  is  used  only  for  clear  solutions,  then  it 
is  of  more  advantage  to  have  it  in  a  horizontal  position.  Fig.  42 
represents  a  horizontal  pressure  tank  designed  to  lift  the  sodium 


FIG.  41. —AIR  BLOW-OFF  DRUM. 
Plan  of  tray. 

hyposulphite  solution.  It  is  made  of  steel  12  ft.  long  and  4  ft. 
6  in.  in  diameter.  The  filling  pipe  P  enters  at  the  head  of  the 
tank  in  order  to  save  grade.  This  tank  has  to  be  placed  in  a  pit 
below  the  bottom  level  of  the  lower  storage  tank;  by  inserting 
the  filling  pipe  at  the  center  of  the  head  a  more  shallow  pit  will 


Steel 
FIG.  42.— HORIZONTAL  PRESSURE  TANK,  FOR  SOLUTION. 

answer,  and  is  more  convenient  for  work.  The  manhole  is  on 
top,  but  near  one  end  of  the  tank.  At  A  the  solution  discharge- 
pipe  is  inserted,  which  extends  nearly  to  the  bottom.  The  part 


202 


HYDROMETALLURGY  OF  SILVER 


of  the  pipe  which  extends  inside  the  tank  is  subject  to  wear  and 
should  be  so  arranged  as  to  permit  of  quick  replacement.  In 
Fig.  43  its  construction  is  shown  in  detail:  A,  a  short  tube,  one 


FIG.  43.  — CAST-IRON  FLANGE  UNION  FOR  DISCHARGE  PIPE  OF 
PRESSURE  TANK. 

The  dimensions  here  given  are  for  2-in.  discharge  pipe  as  used  in  the  vertical 
pressure  tanks.  They  have  to  be  changed  for  other  pressure  tanks  according  to 
size  of  their  respective  discharge  pipe.  The  part  that  extends  into  the  tank  can 
be  easily  removed  and  replaced. 

end  of  which  is  fastened  to  the  tank,  while  the  other  is  provided 
with  a  flange  F,  of  the  shape  shown.  B,  the  pipe  which  extends 
into  the  tank.  The  upper  end  of  it  is  provided  with  the  flange 


TREATMENT  OF  THE  PRECIPITATE  203 

E,  which  is  faced  on  both  sides.  This  flange  fits  loosely  into 
the  recess  of  the  flange  F.  When  the  tube  is  inserted  a  J-in. 
rubber  gasket,  C,  is  placed  on  top  and  covered  with  the  flange  L, 
which  is  attached  to  the  discharge-pipe  M.  This  done,  the 
flanges  L  and  F  are  tightly  drawn  together  by  the  bolts  D,  D. 
The  rubber  gasket  is  strongly  pressed  by  the  bolts  and  makes  a 
perfectly  tight  joint.  By  the  construction  it  can  be  seen  that 
the  pipe  B  can  be  withdrawn  and  replaced  in  a  few  minutes.  All 
pipes  extending  into  a  pressure  tank  ought  to  be  arranged  in  the 
way  just  described. 

Handling  of  the  Sodium  Hyposulphite  Solution.  —  It  was 
stated  above  that  the  solution,  by  precipitating  the  silver  and 
the  other  metals  dissolved  in  it,  is  regenerated,  and  can  be  used 
over  and  over  again  indefinitely.  As  the  solution  works  from 
the  upper  level  down  to  the  lowest  level  of  the  leaching  plant  by 
gravity,  it  has  to  be  elevated  again  in  order  to  keep  it  in  circula- 
tion. The  solution  coming  from  the  precipitation  tanks  is  col- 
lected in  a  number  of  large  but  not  too  deep  tanks  (storage  tanks). 
These  tanks,  usually  four  in  number,  communicate  by  means 
of  iron  pipes  near  the  bottom.  The  solution  coming  from  the 
precipitation  tanks  flows  only  into  the  first  tank  and  into  no 
other,  but  through  the  communicating  pipes  all  four  tanks  are 
filled  simultaneously.  The  fourth  or  last  tank  only  is  connected 
with  the  filling-pipe  of  the  pressure  tank.  This  arrangement  of 
the  four  communicating  tanks  offers  an  excellent  opportunity 
for  settling  any  precipitate  that  may  have  been  drawn  out  in 
decanting  by  carelessness.  Above  the  leaching  tanks  there  is 
another  set  of  four  storage  tanks  of  the  same  dimensions,  arranged 
exactly  as  are  the  lower  tanks.  Into  the  first  of  these  tanks  the 
pressure  tanks  lift  and  discharge  the  solution,  while  from  the 
fourth  one  a  pipe-line  leads  over  all  the  leaching  tanks.  These 
upper  storage  tanks  give  another  opportunity  for  settling.  They 
are  cleaned  once  or  twice  a  year. 

In  large  works  the  circulating  stream  of  solution  is  quite 
voluminous,  and  it  is  advisable  to  have  two  solution  pressure 
tanks,  so  that  a  constant  stream  can  be  maintained. 

Removal  of  Sulphur.  —  The  precipitant  being  a  polysulphide, 
the  precipitate  will  contain  a  large  percentage  of  free  sulphur, 
whether  calcium  or  sodium  sulphide  is  used.  It  is  desirable  to 
remove  this  free  sulphur  from  the  precipitate  before  the  latter 


204 


HYDROMETALLURGY   OF   SILVER 


is  subjected  to  a  final  treatment.  The  best  method  is  to  boil 
it  with  caustic  soda  in  an  iron  tank,  the  caustic  soda  combining 
with  the  free  sulphur  and  forming  sodium  sulphide,  which  serves 
as  precipitant  for  the  silver;  but  care  has  to  be  taken  that  no 
excess  of  caustic  soda  enters  the  stock  solution,  because  it  will 
exercise  a  decomposing  action  on  the  silver  chloride  in  the  ore. 
To  avoid  this  the  sodium  sulphide  solution  thus  obtained  is  con- 
veyed to  the  boiling  vessels  in  which  the  precipitant  is  manufac- 
tured. After  boiling  with  caustic  soda  the  precipitate  shrinks 


FIG.  44. —APPARATUS  FOR  THE  MANUFACTURE  OF  LYE. 
To  be  placed  on  cooling  floor  next  to  roaster  floor. 

much  in  volume  and  becomes  very  heavy.  To  separate  it  from 
the  sodium  sulphide  solution  after  the  main  part  has  been  de- 
canted is  a  very  slow  process  if  done  by  common  filters,  and 
therefore  it  is  much  better  to  use  a  small  filter  press  for  this  pur- 
pose. In  works  where  no  filter  press  is  used,  and  a  common 
filter  has  to  be  employed,  these  badly  filtering  sulphides  can  be 
made  quick  filtering  by  treating  them  with  a  strong  silver  solu- 
tion to  decompose  all  the  sodium  sulphide,  which  is  the  cause  of 
the  bad  filtration.  By  this  method  60  per  cent,  of  the  sulphur 
contained  in  the  precipitate  can  be  regained  and  brought  into 


TREATMENT   OF  THE   PRECIPITATE  205 

a  state  in  which  it  can  be  used  again  as  precipitant,  thus  greatly 
reducing  the  actual  consumption  of  sulphur. 

F.  Sustersic  proposed  to  leach  wood  ashes,  convert  the  lye 
by  boiling  with  caustic  lime  into  caustic  potash  and  boil  the  pre- 
cipitate with  it,  thus  producing  potassium  sulphide,  which  is 
used  as  precipitant.  I  adopted  and  carried  out  this  method. 

In  most  works  wood  is  used  as  fuel  for  the  roasters  and  boilers, 
and  the  ashes  are  thrown  away.  By  adopting  this  method  a 
large  part  of  these  ashes  can  be  utilized.  Fig.  44  illustrates  an 
arrangement  for  the  manufacture  of  lye.  The  tank  T  is  placed 
near  the  roaster  floor,  with  its  rim  just  a  little  below  that  floor, 
so  that  the  wheelbarrow  runway  resting  on  the  rim  of  the  tank 
is  on  a  level  with  the  roaster  floor.  The  tank  is  provided  with  a 
filter  bottom  (not  shown  in  the  figure).  The  ashes  from  the 
roasters  and  also  from  the  boilers,  if  the  latter  are  conveniently 
situated,  are  wheeled  and  dumped  into  the  tank,  which  is  filled 
one-third  full  with  water.  This  is  done  to  protect  the  tank  and 
filter  from  the  hot  ashes,  which  contain  many  small  pieces  of  burn- 
ing coal.  When  the  tank  is  filled,  water  is  allowed  to  flow  on  top 
of  the  ashes.  The  outflowing  lye  is  collected  in  the  iron  tank  R. 
When  the  lye  begins  to  get  weak  the  hose  H  is  placed  in  the 
trough  L,  which  leads  outside  the  building,  and  the  charge  is 
allowed  to  drain  and  then  replaced  by  fresh  ashes.  From  the 
iron  tank  R  a  pipe-line  leads  to  the  calcium  sulphide  boilers, 
which  can  also  be  used  for  the  manufacture  of  caustic  potash. 
One  of  the  boilers  is  charged  with  lye,  to  which  milk  of  lime  is 
added.  The  mixture  is  then  boiled.  Samples  are  taken  from 
time  to  time,  filtered,  and  to  the  filtrate  a  few  drops  of  hydro- 
chloric acid  added.  If  this  causes  effervescence  boiling  is  con- 
tinued, but  if  after  a  while  effervescence  is  still  caused  by  the 
acid,  then  more  milk  of  lime  has  to  be  added.  W^hen  finished 
the  charge  is  pressed  through  the  same  filter  press  which  is  used 
for  calcium  sulphide.  The  filtrate  is  collected  in  an  iron  storage 
tank  placed  alongside  of  the  calcium  sulphide  tank. 

Fig.  45  represents  a  system  of  three  pressure  tanks;  they 
serve  for  the  treatment  of  the  precipitate,  and  are  placed  in  the 
refinery  of  the  works.  The  pressure  tank  A  is  charged  with 
precipitate  from  the  precipitation  tanks  through  the  pipe  D. 
From  there  it  is  forced  through  the  pipe  E  into  a  filter  press. 
When  pressed,  the  precipitate  is  charged  into  pressure  tank  C, 


206 


HYDROMETALLURGY   OF  SILVER 


through  the  opening  in  the  manhole.  Caustic  potash  solution 
is  added,  which  is  done  by  opening  the  valve  of  pipe  G,  which 
latter  is  connected  with  the  caustic  potash  storage  tank.  The 
mixture  is  boiled,  and  when  finished  forced,  through  pipe  H,  into 
a  second  filter  press.  The  filtrate  from  this  press,  which  is  potas- 
sium sulphide,  flows  into  pressure  tank  B,  through  pipe  K,  whence 
it  is  lifted  up  through  the  pipe  M  and  charged  into  the  calcium 
sulphide  boiling  tank. 


FIG.  45. —PRESSURE  TANKS  FOR  TREATMENT  OF 
PRECIPITATE. 

As  by  this  operation  about  60  per  cent,  of  the  sulphur  is  re- 
gained in  a  shape  in  which  it  can  be  directly  applied  as  precipitant, 
and  at  the  same  time  the  precipitate  is  freed  from  its  surplus  sul- 
phur, and  as,  besides,  the  caustic  potash  does  not  cost  more  than 
the  labor  and  the  lime,  this  is  surely  a  very  economical  operation. 


TREATMENT  OF  THE  PRECIPITATE  207 

Another  way  of  expelling  and  regaining  the  free  sulphur 
from  the  precipitate  is  by  distillation.  The  moist  precipitate  is 
charged  into  retorts  and  heated,  the  sulphur  vapors  being  con- 
ducted to  brick  chambers  and  condensed  as  flowers  of  sulphur. 
I  used  this  method  years  ago  on  a  large  scale  with  satisfactory 
results  as  to  the  amount  of  sulphur  regained,  but  the  cast-iron 
retorts  did  not  last  long  enough,  principally  on  account  of  care- 
lessness on  the  part  of  the  men  in  charge,  who  overheated  them; 
and  as  transportation  of  such  heavy  castings  into  the  mountains 
was  very  difficult  and  expensive,  this  method  was  discarded. 
However,  with  careful  firing,  and  in  localities  where  transporta- 
tion facilities  are  better,  it  can  be  applied  to  great  advantage. 
While  the  extraction  of  the  free  sulphur  will  not  be  so  complete  as 
by  boiling  the  precipitate  with  caustic  soda  or  potash,  the  opera- 
tions are  fewer.  The  product  is  dry  and  ready  immediately  for 
further  treatment,  while  in  the  former  method  the  product  has 
to  undergo  the  processes  of  filtering,  washing  and  drying. 

A  third  method,  by  which,  however,  the  sulphur  is  lost,  con- 
sists in  burning  it  off  in  a  small  reverberatory  furnace.  An  actual 
roasting  is  not  required;  in  fact  it  ought  to  be  avoided,  to  prevent 
loss  by  volatilization.  The  sulphides  ought  to  be  charged  dry, 
but  if  they  are  charged  moist  they  should  remain  undisturbed  in 
the  moderately  heated  furnace  until  dry,  to  avoid  the  generation 
of  rapidly  evolving  steam,  which  is  apt  to  carry  away  fine  par- 
ticles of  the  already  dry  part  of  the  precipitate,  thus  causing  a 
loss.  Even  if  the  precipitate  has  been  previously  dried  in  special 
ovens,  the  heat  in  the  beginning  has  to  be  kept  very  low  for  some 
time,  to  avoid  mechanical  loss  by  steam,  because  only  seldom 
will  the  precipitate  be  perfectly  free  from  moisture  after  leav- 
ing the  drying  ovens.  Later  the  temperature  is  increased  to  ignite 
the  sulphides.  They  commence  to  burn  with  a  blue  flame  near 
the  fire-bridge,  and  the  flame  spreads  gradually  over  the  whole 
charge.  When  this  takes  place  the  fire  has  to  be  lowered  to 
avoid  overheating.  No  stirring  should  be  done  until  the  flame 
ceases;  then  a  gentle  stirring  is  given.  This  brings  up  new  flames, 
which,  however,  do  not  last  long,  but  reappear  if  the  charge  is 
stirred  again.  Stirring  is  repeated  until  the  flame  ceases  entirely. 
The  temperature  has  to  be  kept  so  that  when  the  flame  ceases 
the  charge  is  perfectly  dark,  which  indicates  that  no  actual  roast- 
ing of  the  material  took  place  and  that  only  the  free  sulphur  was 


208 


HYDROMETALLURGY   OF  SILVER 


burned  off.  If  the  precipitate  is  treated  in  this  way  no  loss  by 
volatilization  will  take  place,  and  that  is  all  which  is  required, 
because  for  the  further  treatment  of  the  precipitate  an  actual 
roasting,  a  changing  of  the  sulphides  to  oxides  and  sulphates,  is 
not  necessary;  in  fact,  is  hurtful.  Mr.  Stetefeldt,  in  his  book  on 
lixiviation,  stated  the  loss  to  be  as  high  as  6  and  12  per  cent.,  but 
this  is  not  so.1  However,  there  is  a  chance  of  loss  by  this  method 
if  it  is  not  properly  executed,  especially  if  the  precipitate  con- 
tains antimony  or  is  charged  wet  into  the  furnace. 


T1 


FIG.  46.  — DRYING  AND  ROASTING  FURNACE  FOR  SILVER 
PRECIPITATE. 

Fig.  46  represents  a  vertical  and  Fig.  47  a  horizontal  section 
of  a  small  reverberatory  furnace  for  burning  the  silver  precipitate. 
Where  the  boiling  of  the  precipitate  with  caustic  potash  or  soda 
is  adopted  this  furnace  can  be  used  for  drying  the  treated  pre- 
cipitate. 

The  first  method,  i.e.,  boiling  with  caustic  soda  or  potash,  is 
the  most  rational  and  entirely  excludes  any  loss  of  silver,  and 
offers  the  additional  advantage  of  regaining  about  60  per  cent, 
of  the  sulphur. 

Refining  the  Precipitate.  —  The  refining  is  done  with  litharge 

1  Stetefeldt  based  his  statements  on  reports  received  from  Cusihuiriachic, 
while  the  works  were  under  the  management  of  Mr.  Dagget.  Later  it 
developed  that  the  great  loss  attributed  to  the  burning  of  the  sulphides  was 
caused  by  the  dishonesty  of  the  man  in  charge  of  that  part  of  the  process, 
who  -stole  systematically  part  of  each  charge. 


TREATMENT   OF  THE   PRECIPITATE 


209 


on  a  lead  bath  in  the  cupeling  furnace.     Other  methods   have 
been  tried,  but  so  far  not  with  much  success. 

If  the  refining  is  done  on  the  lead  bath  usually  English  cupel- 
ing furnaces  are  used,  which  are  so  constructed  that  the  test  can 
be  dipped  toward  the  front  to  pour  the  refined  silver  into  the 
molds.  If,  however,  the  works  produce  larger  quantities  of  precip- 
itate, it  is  more  advantageous  to  cupel  in  a  larger  furnace,  simi- 
lar in  construction  to  the  German  cupeling  furnace,  with  the 
exception  that  the  bottom  is  not  stamped  direct  into  the  circular 
space  left  in  the  brickwork,  but  into  a  circular  cast-iron  pan,  the 
bottom  of  which  is  perforated  with  J-in.  holes.  The  pan  is 


FIG.  47.— DRYING  AND  ROASTING  FURNACE  FOR  SILVER 
PRECIPITATE. 

made  in  four  sections,  which  are  kept  together  with  a  few  bolts, 
which  construction  permits  the  pan  to  expand  without  breaking. 
This  pan  sets  above  a  second  pan,  2  in.  larger  in  diameter 
and  only  2  in.  deep,  and  rests  on  bricks  set  in  the  lower  pan  on 
their  4-in.  side.  This  flat  pan  serves  as  a  guard  against  loss  of 
silver,  in  case  it  should  happen  that  the  bottom  of  the  test  cracks. 
The  space  underneath  communicates  with  the  outside  of  the  fur- 
nace by  four  channels,  through  which  air  circulates  and  cools 
the  bottom,  while  in  drying  the  test  they  serve  as  vents  for  the 
vapors.  The  perforated  pan  is  5J  to  6  ft.  in  diameter  and  12  in. 
deep,  and  has  a  square  cut  12  in.  wide  and  10  in.  deep  for  the 
litharge  bridge.  This  cut  is  placed  toward  the  front.  To  the 


210  HYDROMETALLURGY  OF  SILVER 

left  of  it  is  the  fireplace  and  to  the  right  the  flue,  which  is  pro- 
vided with  a  damper,  while  the  back  of  the  furnace  is  arranged 
to  admit  the  blast.  The  furnace  proper  is  covered  with  a  dome 
made  of  boiler  iron  and  lined  with  clay.  This  dome  is  lifted  and 
can  be  swung  to  one  side  by  a  crane. 

A  cheap  and  excellent  material  for  the  cupel  is  a  mixture  of 
one  part  of  clay  (by  volume)  and  three  parts  of  lime  rock,  pul- 
verized and  well  mixed.  Both  ingredients  must  be  free  from 
quartz  and  ore  particles,  for  which  reason,  if  the  crushing  has  to 
be  done  by  machinery  that  is  also  used  for  pulverizing  ore,  such 
machinery  must  first  be  cleaned  very  carefully. 

After  the  material  is  well  mixed  part  of  it  is  spread  on  a  clean 
floor,  sprinkled  with  water,  and  quickly  worked  with  shovels  so 
that  the  mixture  becomes  uniformly  moist.  The  mixture  should 
not  be  made  too  moist,  as  otherwise  in  refining  the  bottom  is  apt 
to  come  up.  The  material  is  in  proper  condition  if  a  handful  of 
it,  squeezed  hard,  forms  a  ball,  which  may  be  handled  gently,  but 
should  crumble  into  its  former  condition  by  a  slight  pressure  with 
the  fingers.  In  preparing  the  test  the  perforated  pan  is  filled 
about  6  in.  with  the  prepared  material,  then  stamped  down  with 
iron  bars. 

These  bars  are  made  of  a  piece  of  round  iron  1 J  in.  in  diameter 
and  8  to  10  in.  long,  one  end  of  which  has  the  shape  of  an  egg 
while  the  other  is  welded  to  a  IJ-in.  gas-pipe  5  ft.  long.  The 
material  in  the  pan  is  leveled  and  beaten  in  by  two  men  stand- 
ing on  boards  laid  across  the  brickwork.  They  commence  to 
beat  in  the  center,  pursuing  a  spiral  course,  stamping  with  the 
egg  point  in  a  perpendicular  direction,  by  a  lift  of  about  eight 
inches,  striking  with  the  rod  close  to  each  preceding  stroke. 
When  by  a  screw-like  advance  the  stamping  has  reached  the  side 
of  the  pan,  it  has  to  be  carried  on  back  to  the  center  in  the  same 
way,  then  again  to  the  side  and  so  on,  till  about  two  inches  of 
loose  material  remains. 

If  a  hole  can  be  scratched  easily  with  the  finger  in  the 
stamped  mass,  the  bar  must  be  used  with  more  force.  Care 
must  be  taken  always  to  have  still  two  inches  of  loose  material 
above  the  stamped  mass  when  a  new  charge  is  put  in,  because 
if  the  whole  is  beaten  hard,  the  next  charge  will  not  unite  per- 
fectly with  the  under  layer.  The  beating  on  the  second  charge 
is  done  in  the  same  way,  and  so  on  until  the  hard-beaten  mass 


TREATMENT  OF  THE  PRECIPITATE  211 

reaches  about  two  inches  below  the  rim  of  the  pan.  Then  a 
6-in.  high  iron  ring  fitting  the  rim  of  the  pan  is  placed  on  the  rim,  a 
new  charge  added,  and  stamping  continued  until  the  hard  stamped 
mass  reaches  about  three  inches  above  the  rim  of  the  pan. 
Then  the  loose  material  is  removed,  the  ring  pulled  up,  and  the 
hard  mass  is  cut  down  roughly  by  means  of  a  hatchet  or  an  adze 
to  about  an  inch  above  the  rim  of  the  pan,  and  then  leveled  with 
a  sharp  short-handled  scraper.  This  done,  a  circle  is  scratched 
in,  leaving  a  margin  of  about  four  inches  around  the  periphery 
of  the  pan.  Inside  this  circle  the  hard  stamped  mass  is  cut  out 
spherically  with  a  short-handled  adze  about  7  to  8  in.  deep;  then 
made  smooth  by  a  scraper  with  a  curved  blade.  When  the 
bottom  is  finished  the  hood  is  put  over  it  by  means  of  the  crane 
and  a  gentle  fire  started  for  drying.  It  is  well  to  scatter  a  thin 
layer  of  ashes  over  the  surface.  The  fire  should  be  kept  about 
twelve  hours,  then  the  ashes  removed  and  the  lead  charged. 
When  the  lead  is  melted  and  becomes  red  hot  some  coarsely 
pulverized  litharge  is  charged,  enough  to  cover  the  whole  surface 
of  the  bath.  A  strong  fire  is  kept  and  the  precipitate,  mixed  with 
litharge,  is  put  on  the  bath  in  small  charges.  When  this  is  going 
on  the  draft  has  to  be  checked  by  the  damper  so  that  the  flames 
will  come  out  through  the  working  door.  This  precaution  is 
taken  to  reduce  the  amount  of  such  valuable  dust  to  be  carried 
out  by  the  draft  into  the  dust-chambers.  After  the  charge  is 
introduced  the  damper  is  opened  again,  and  when  the  charge  is 
melted  the  upper  layer  is  worked  gently  with  a  hoe,  and  a  new 
charge  is  introduced.  This  is  repeated  until  all  the  precipitate  is 
charged.  Then  some  more  litharge  is  added,  the  working  door 
closed  and  the  strong  fire  continued.  It  will  be  noticed  that, 
when  everything  is  well  melted,  around  the  periphery  of  the 
bath  many  small  bubbles  will  appear.  This  is  caused  by  the 
carbonic  acid  which  the  lime  rock  gives  off,  and  does  not  injure 
the  bottom. 

When  the  last  charge  is  well  melted  the  top  layer  is  drawn  off 
over  the  litharge  bridge  by  means  of  a  hoe  until  the  surface  be- 
comes bright.  Then  the  blast  is  turned  on,  playing  on  the  sur- 
face, which  makes  the  bath  fume  profusely.  This  is  caused  by 
the  oxidation  of  the  lead  matte  which  had  formed  during  melting 
and  which  lays  on  top  of  the  lead.  These  fumes  smell  strongly 
of  sulphurous  acid.  After  two  or  three  hours  the  surface  loses 


212 


HYDROMETALLURGY   OF  SILVER 


...A. 


TREATMENT  OF  THE  PRECIPITATE  213 

its  brightness,  a  yellow-red  ring  of  litharge  forms  around  the 
periphery,  and  little  islands  of  the  same  color  are  formed  where 
the  blast  strikes.  They  float  about  until  they  reach  the  ring 
of  litharge  around  the  periphery  and  unite  with  it.  The  width  of 
this  ring  increases  until  nearly  the  whole  surface  of  the  bath  is 
covered.  When  this  has  happened  a  flat  channel  is  cut  into  the 
litharge  bridge,  through  which  the  litharge  flows  off.  The  flow 
is  to  be  regulated  so  that  the  ring  is  kept  about  8  in.  wide.  Care 
is  to  be  taken  that  no  metal  flows  over  the  bridge.  The  tempera- 
ture has  to  be  kept  so  that  the  litharge  runs  freely  over  the  bridge 
but  cuts  the  channel  very  little.  If  the  heat  is  too  high  the 
litharge  will  cut  the  channel  rapidly  deeper  and  metal  will  flow 
out.  On  the  other  hand,  if  the  temperature  is  too  low  the  litharge 
will  flow  sluggishly  over  the  bridge  and  form  a  soft  crust  on  the 
inside,  near  the  channel.  Too  low  a  temperature  has  to  be 
carefully  avoided,  as  the  whole  bath  may  freeze,  especially  to- 
ward the  end  of  the  process.  Likewise  should  a  too  high  tem- 
perature be  avoided,  because  then  the  litharge  cuts  around  the 
periphery  of  the  bottom,  thus  ruining  it,  and  eats  out  the  channel 
too  quickly,  besides  which  more  silver  and  lead  are  volatilized. 

At  the  end  of  the  operation,  when  less  litharge  is  formed, 
the  bath  becomes  covered  with  a  net-like  coat,  moving  on  the 
convex  surface  and  consisting  of  litharge,  between  which  the 
silver  glances  through  in  spots.  These  spots  grow  larger,  till  at 
last  the  net  breaks  and  the  litharge  slides  to  the  sides,  producing 
a  display  of  colors.  Before  the  end,  but  at  the  time  when  the 
formation  of  litharge  becomes  scanty,  the  iron  molds  are  arranged 
on  an  iron  bench  set  in  front  of  the  furnace.  A  long-handled  iron 
ladle  hooked  to  a  chain  reaching  down  from  above  is  heated  by 
placing  it  to  one  side  inside  the  furnace.  As  soon  as  the  surface 
of  the  silver  turns  bright  it  is  ladled  into  the  molds.  During 
ladling  the  temperature  has  to  be  kept  high. 

Should  the  litharge  become  stiff  and  sluggish  toward  the  end, 
even  if  the  temperature  is  increased,  it  is  a  sign  that  there  is  not 
enough  lead  in  the  bath  to  separate  the  copper,  and  without  loss 
of  time  some  lead  has  to  be  added,  but  on  the  side  of  the  test 
and  not  into  the  bath,  because  this  would  cool  it  too  much  and 
might  freeze  it. 

All  the  products  which  are  formed  in  cupeling  contain  silver. 
The  first  slag  drawn,  as  well  as  the  last  litharge  near  the  end  of 


214  HYDROMETALLURGY  OF  SILVER 

the  operation,  is  the  richest,  while  the  poorest  and  purest  is 
obtained  during  the  middle  of  the  operation.  The  latter  goes 
back  to  the  process  at  the  next  cupellation.  The  products  that 
are  too  rich  it  is  best  to  sell  to  the  smelting  works,  as  it  would  be 
too  expensive  to  treat  them  by  themselves. 

As  it  cannot  be  avoided  that  some  dust  of  the  precipitate, 
which  is  very  valuable,  is  carried  away  by  the  draft  during  charg- 
ing, and  as  the  lead  fumes  from  the  bath  also  contain  silver,  it  is 
very  necessary  to  provide  for  an  effective  dust  collector.  Fig.  48 
illustrates  a  four-shaft  system  of  O.  Hofmann's  dust  collector. 
It  is  inserted  in  the  main  flue  of  the  refinery,  so  that  all  the  fumes 
and  dust  from  the  different  furnaces  have  to  pass  through  it. 


XVI 

CONSTRUCTION   OF  TROUGHS 

In  lixiviating  works,  troughs  are  extensively  used  for  con- 
veying the  base  metal  and  the  silver  solution,  and  it  is  of  great 
importance  to  have  and  keep  them  perfectly  tight.  The  solu- 
tions will  corrode  metals,  including  lead,  and  the  troughs  there- 
fore have  to  be  made  of  wood,  but  it  is  a  very  difficult  task  to 
make  and  keep  troughs  perfectly  tight,  especially  if  they  have 
only  a  slight  inclination  and  the  solutions  do  not  move  swiftly. 
Leakage  occurs  principally  in  places  where  the  boards  are  spliced. 
White  lead,  tar,  pitch,  putty,  etc.,  are  of  but  little  avail.  White 
lead  is  the  worst,  and  ought  never  to  be  used  by  a  carpenter  to 
make  a  joint  waterproof.  As  long  as  the  paint  is  wet  it  will 
be  tight,  but  when  the  wood  absorbs  the  oil  and  the  paint  dries, 
it  contracts  into  numerous  threads  and  wrinkles  and  hardens, 
forming  minute  channels  through  which  the  solution  will  find 
its  way.  The  places  where  the  white  lead  was  applied  will  not 
swell,  and  the  trough  will  be  much  less  tight  than  if  no  cement 
had  been  used  at  all.  In  order  to  convey  the  solution  to  the 
different  tanks,  branch  troughs  have  to  be  used.  One  end  of 
these  troughs  is  placed  under  the  main  trough,  and  right  above 
it  one  or  more  holes  are  bored  into  the  bottom  of  the  main  trough. 
These  holes  are  closed  by  long  wooden  plugs.  The  loosening  and 
tightening  of  these  plugs  is  done  by  mallets,  and  therefore  the 
main  trough  is  subjected  to  considerable  rough  usage,  and  the 
pitch  or  putty  will  crack  and  start  leakage. 

In  running  the  large  blue  vitriol  plant  of  the  Consolidated 
Kansas  City  Smelting  and  Refining  Company  of  Argentine, 
Kansas,  which  I  designed  and  erected,  I  was  very  much  annoyed 
by  trough  leakage.  There  were  over  2000  ft.  of  troughs  in  use. 
To  guard  against  leakage  all  the  troughs  were  lined  with  6-lb. 

215 


216  HYDROMETALLURGY  OF  SILVER 

sheet  lead  properly  put  in  by  experienced  lead  burners,  but  be- 
fore a  year  passed  the  troughs  began  to  leak  in  different  places. 
Searching  for  the  reason  it  was  found  that  this  leakage  was 
caused  by  a  peculiar  property  of  the  lead.  The  copper  solution 
passing  through  the  troughs  was  hot.  It  did  not  flow  continually, 
but  in  charges,  so  that  the  lead  was  exposed  alternately  to  the 
hot  liquor  and  to  the  cooling  effect  of  the  air.  The  lead  lining, 
which  originally  was  perfectly  smooth,  was  found  to  be  full  of 
wrinkles.  These  wrinkles  could  have  been  originated  only  by  the 
expansion  of  the  lead.  By  special  experiments  I  convinced  my- 
self that  lead  is  a  metal  which  when  heated  expands,  but  when 
cooled  does  not  contract  as  much  as  it  had  expanded;  and  if  a 
sheet  of  lead  is  alternately  heated  and  cooled  it  continues  to  grow 
larger  and  larger.  If  there  is  no  room  for  free  expansion,  as  in  a 
trough  or  tank,  the  sheet  has  to  fold  up  in  wrinkles.  If  this  is 
continued,  the  sheet  gets  so  thin  in  certain  places  that  it  finally 
breaks. 

For  some  time  the  leakage  was  stopped  by  burning  a  piece 
of  sheet  lead  over  the  leaking  places,  but  before  long  the  leaks 
became  so  frequent  that  the  item  of  trough  repairing  became 
seriously  high,  and  it  was  concluded  to  renew  all  the  troughs. 
Lead  lining  was  of  course  discarded.  Previous  experiments 
with  different  kinds  of  soft  wood  showed  that  California  redwood 
resisted  best  and  longest  the  action  of  a  hot  copper  solution. 
Other  experiments  were  made  to  prepare  a  cement  which  softened 
but  did  not  melt  at  a  temperature  of  90  to  96  deg.  C.,  and  which 
in  cooling  did  not  become  hard  and  brittle  but  remained  pliable. 
Such  a  cement  was  found  by  boiling  lard  oil  with  rosin  and  rubber 
and  red  oxide  of  iron.  Waste  pieces  of  sheet  rubber  and  old 
pieces  of  rubber  belting  were  used  to  supply  the  rubber.  The 
fibers  of  the  belting  were  removed  with  iron  hooks  and  forks. 
The  oxide  of  iron  was  added  last.  The  boiling  was  done  in  an 
iron  kettle  with  a  fireplace  below.  The  cement  or  paste  was 
applied  hot.  Fig.  49  shows  the  construction  of  the  trough  in 
cross-section.  The  joints  of  sides  and  bottom  are  made  step- 
shaped,  as  shown  in  the  drawing,  and  are  first  coated  with  the 
hot  cement,  then  screwed  tight  with  brass  screws.  The  corners 
are  filled  with  triangular  wooden  moldings,  which  are  first  well 
coated  with  cement.  They  are  kept  in  place  and  tightly  drawn  up 
by  brass  screws.  If  the  corners  are  first  well  cemented,  covered 


CONSTRUCTION   OF  TROUGHS 


217 


by  a  strip  of  thin  sheet  rubber,  and  then  the  molding  screwed  on, 
additional  security  for  tight  joints  is  obtained.  The  troughs  are 
made  in  sections  of  the  length  of  the  boards,  and  both  ends  cut 


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Cast  Iron 

FIG.  51. -TROUGH  UNION. 


square.     The  ends  of  two  sections  are  brought  together,  a  pure 
rubber  gasket,  made  to  order  in  one  piece,  J  or  f  in.  thick,  is 


218  HYDROMETALLURGY   OF  SILVER 

placed  between  them  and  drawn  tight  by  bolts.  Fig.  50,  a  side 
view,  illustrates  the  manner  in  which  the  two  ends  of  the  trough 
are  drawn  together.  C  and  D  are  flanged  iron  castings  of  the 
shape  of  the  cross-section  of  the  trough,  but  smaller,  so  that  they 
will  fit  into  a  groove  2J  in.  wide  and  i  in.  deep  cut  into  the  three 
sides  of  the  trough  and  about  6  in.  from  the  end.  These  castings 
are  made  in  pairs — one  right,  one  left,  as  shown  in  Fig.  51.  When 
placed,  they  are  fastened  to  the  trough  with  wood  screws,  and 
by  means  of  the  bolts  E  the  ends  of  two  troughs  are  pressed  to- 
gether. Thus  perfectly  tight  joints  of  the  different  sections  are 
obtained.  At  distances  of  four  feet  wooden  pieces  are  nailed 
across  the  trough  to  prevent  the  sides  from  spreading. 

Troughs  made  in  this  way  are  perfectly  tight  and  will  stay 
so  for  years. 


XVII 

TROUGH    LIXIVIATION 

IN  tank  lixiviation,  the  extraction  of  the  silver  from  chlori- 
dized  ore  by  solutions  of  hyposulphite  salts  is  performed  by  fil- 
tration. The  ore  particles  are  kept  stationary,  while  the  solvent 
moves  down  through  the  mass  of  ore.  The  quickness  of  extrac- 
tion, other  conditions  alike,  is  in  direct  proportion  to  the  rapidity 
of  the  movement  of  the  solvent  through  the  ore.  The  solution, 
if  left  in  contact  with  the  ore  without  moving,  displays  but  very 
little  dissolving  energy.  If  the  filtration  is  interrupted  for  ten  or 
twelve  hours,  and  thus  solution  and  ore  are  left  in  complete  con- 
tact for  that  length  of  time,  it  will  be  found  that,  when  filtration 
is  started  again,  the  outflowing  solution  is  but  very  little  more 
saturated  with  silver  than  it  was  at  the  time  of  interruption, 
and  that  the  ten  or  twelve  hours  were  almost  a  complete  loss  in 
the  total  time  of  extraction.  Notwithstanding  the  long  contact, 
the  solution  had  not  become  saturated  with  metal  chlorides  to 
its  full  dissolving  capacity.  A  rapid  movement  of  the  solvent 
through  the  ore  is  essential  to  a  quick  extraction.  This  fact  is 
well  known;  and  the  endeavor  of  leachers  has  been  to  hasten 
extraction  by  increasing  the  rate  of  filtration.  Siphons,  vacuum 
pumps  and  other  devices  have  been  used  with  more  or  less  success, 
but  none  of  them  has  given  full  satisfaction. 

I  have  found  that,  if  chloridized  ore,  after  the  base-metal 
chlorides  are  removed,  is  brought  into  rapid  contact  with  a 
proper  volume  of  moving  sodium  hyposulphite  solution,  the 
silver  chloride  contained  in  the  ore  dissolves  almost  instantly, 
and  that  it  is  rather  the  volume  of  the  solvent  than  its  concen- 
trated state  which  produces  this  effect.  Such  favorable  con- 
ditions cannot  be  attained  in  tanks.  The  rapidity  with  which 
a  certain  volume  of  the  solvent  can  be  brought  into  contact 
with  the  ore  particles  is  limited  by  the  speed  with  which  the 

219 


220  HYDROMETALLURGY   OF  SILVER 

solution  descends  through  the  ore;  and  thus  the  leaching  time 
in  tanks  cannot  be  shortened  beyond  the  limit  set  by  the  filter- 
ing capacity  of  the  ore.  In  a  trough,  however,  these  favorable 
conditions  can  be  attained  by  gradually  introducing  the  ore  into 
the  moving  stream  of  the  solvent.  The  ore  can  thus  be  brought 
into  rapid  contact  with  any  desired  quantity  of  the  solvent,  and 
moves  in  and  with  the  stream.  The  effect  is  astonishing.  Ore 
charged  at  the  upper  end  of  a  trough  not  longer  than  12  to  15  ft. 
will  leave  the  trough  at  the  lower  end  as  tailings,  having  yielded 
all  its  silver  chloride  to  the  solvent.  This  is  accomplished  dur- 
ing the  very  short  time  of  4.7  seconds  which  it  takes  the  pulp  to 
rush  through  the  trough.  But  in  order  to  obtain  satisfactory 
results,  that  is,  to  extract  all  the  silver  chloride  contained  in  the 
ore  as  shown  by  the  chlorination  test  assay,  it  is  necessary,  and 
of  great  importance,  to  maintain  a  certain  proportion  of  solvent 
and  ore,  which  proportion  depends  on  the  nature  of  the  ore.  By 
numerous  experiments  I  have  found  that  all  kinds  of  silver  ores, 
no  matter  how  differently  they  behave  in  tank  lixiviation  with 
regard  to  the  length  of  time  required  for  the  extraction  of  the 
silver,  will  yield  their  silver  chloride  in  the  same  short  time. 
They  behave  all  alike  in  this  respect,  but  only  if  the  proper  pro- 
portion between  solvent  and  ore  which  each  respective  ore  re- 
quires is  maintained.  It  is  interesting  to  observe  that,  when  two 
ores  of  different  chemical  character  but  of  equal  filtering  property 
are  treated  in  the  trough,  the  one  which  in  tank  lixiviation  re- 
quires the  longer  time  to  yield  its  silver  chloride  to  the  solvent 
will  need  in  the  trough  a  larger  volume  of  solvent  than  the  ore 
which  requires  in  the  tank  less  time  for  the  extraction  of  the 
silver. 

Lead-bearing  ores  require  a  long  leaching  time,  for  the  reason 
that  lead  sulphate  reduces  greatly  the  dissolving  energy  of  sodium 
hyposulphite  for  silver.  In  the  ordinary  lixiviation  the  solution 
becomes  more  saturated  with  lead  sulphate  as  it  descends  through 
the  ore  and  loses  proportionally  its  dissolving  energy.  As  the 
solubility  of  the  lead  sulphate  increases  with  the  concentration 
of  the  sodium  hyposulphite  solution,  a  stronger  solution  does  not 
hasten  the  process;  but  if  we  bring  the  ore  rapidly  in  contact 
with  a  large  volume  of  hyposulphite  solution,  the  latter  retains 
enough  of  its  dissolving  energy  to  produce  a  quick  silver  extrac- 
tion. The  presence  of  lead  sulphate,  therefore,  does  not  retard 


TROUGH  LIXIVIATION  221 

trough  lixiviation;  it  merely  entails  the  use  of  larger  quantities 
of  solvent. 

Results  of  much  interest,  obtained  by  me  in  experimenting 
with  ore  from  the  Cusihuiriachic  Mining  Company,  Chihuahua, 
Mexico,  illustrate  the  importance  of  maintaining  a  certain  pro- 
portion of  solution  and  ore  to  obtain  satisfactory  results.  The 
ore  contains  considerable  lead,  and  the  extraction  by  common 
tank  lixiviation  required  on  an  average  nine  hours  for  base- 
metal  and  53.8  hours  for  silver  leaching,  in  all  62.8  hours.  It 
was  roasted  in  Howell  furnaces.  For  this  experiment  a  1.6  per 
cent,  solution  was  used.  The  roasted  ore  was  first  leached  with 
water  to  remove  the  base-metal  chlorides  before  treating  it  in 
the  trough. 

The  roasted  ore  contained   27.12  oz.  silver  per  ton. 

The  chlorination  test  tailings  called  for  .   3.94  oz.  silver  per  ton. 
The  ore  and  solution  passed  through  43  feet  of  trough. 

PROPORTION  OF  SOLUTION  AND  ORE  IN  WEIGHTS 

6   weights   solution   to    1    ore,  the   trough   tailings   contained 

14.58  oz.  silver  per  ton. 

8  weights  solution  to  1  ore,  the  trough  tailings  contained 

6.56  oz.  silver  per  ton. 

12  weights  solution  to  1  ore,  the  trough  tailings  contained 

5.25  oz.  silver  per  ton. 

18  weights  solution  to  1  ore,  the  trough  tailings  contained 

4.37  oz.  silver  per  ton. 

24  weights  solution  to  1  ore,  the  trough  tailings  contained 

4.37  oz.  silver  per  ton. 

As  the  length  of  the  troughs  was  in  all  cases  alike,  the  time 
of  contact  of  solution  and  ore  was  the  same,  and  as  the  strength 
of  the  solution  was  also  similar,  the  difference  in  the  extraction 
was  caused  only  by  the  volume  of  solution.  The  results  show 
that  the  extraction  is  in  direct  proportion  to  the  volume  of  solu- 
tion until  the  maximum  is  reached,  when  an  increase  of  solution 
does  not  improve  the  extraction  any  more.  The  results  further- 
more show  that  for  the  Cusihuiriachic  ore  the  proportion  is  18 
weights  of  solution  to  one  of  ore,  probably  less,  but  more  than 
12  to  1.  In  this  experiment  the  trough  tailings  contained  0.43  oz. 
per  ton  more  silver  than  the  chlorination  test  tailings  called  for. 
This  shortage  in  extraction  was  caused  by  the  fact  that  the  1.6 
per  cent,  solution  was  too  strong,  as  numerous  subsequent  ex- 
periments demonstrated.  A  solution  containing  but  0.5  per  cent, 
sodium  hyposulphite  gives,  as  in  tank  lixiviation,  the  best  results. 


222  HYDROMETALLURGY  OF  SILVER 

The  idea  suggested  itself  to  investigate  the  behavior  of  the 
base-metal  chlorides  contained  in  the  roasted  ore  in  a  moving 
stream  of  water,  and  it  was  found  that  they  dissolved  just  as 
rapidly  as  the  silver  chloride  in  the  sodium  hyposulphite  solution, 
and  that  passing  through  the  same  short  length  of  trough  the 
solution  was  accomplished,  with  the  exception  of  a  certain  per- 
centage of  sodium  sulphate,  which,  however,  the  ore  is  always 
found  to  contain  in  tank  lixiviation  at  the  time  when  base-metal 
leaching  is  stopped  and  silver  leaching  commenced. 

We  have  learned  in  tank  lixiviation  that,  by  leaching  the  ore 
charge  with  water  in  order  to  remove  the  soluble  metal  chlorides, 
the  water  becomes  so  charged  with  these  chlorides  that  it  dissolves 
silver  chloride  like  sodium  chloride  does,  and  that  the  solubility 
of  the  silver  chloride  increases  with  the  concentration  of  the  solu- 
tion. To  investigate  this  important  feature  of  the  lixiviation 
process,  I  made  the  following  laboratory  test:  Freshly  prepared 
silver  chloride  was  introduced  into  solutions  of  sodium  chloride 
which  had  a  temperature  of  120  deg.  F.,  and  was  left  in  contact 
for  five  minutes,  during  which  time  the  solution  was  vigorously 
agitated,  and  then  filtered.  The  filtrate  was  tested  for  silver 
with  sodium  polysulphide,  and  it  was  found: 

Solution  of  2  per  cent,  sodium  chloride  at  120  deg.  F.  dissolved  in 

5  minutes  no  silver. 

Solution  of  3  per  cent,  sodium  chloride  at  120  deg.  F.  dissolved  in 

5  minutes  no  silver. 

Solution  of  4  per  cent,  sodium  chloride  at  120  deg.  F.  dissolved  in 

5  minutes  no  silver. 

Solution  of  5  per  cent,  sodium  chloride  at  120  deg.  F.  dissolved  in 

5  minutes  no  silver. 

Solution  of  6  per  cent,  sodium  chloride  at  120  deg.  F.  dissolved  in 

5  minutes no  silver. 

Solution  of  7  per  cent,  sodium  chloride  at  120  deg.  F.  dissolved  in 

5  minutes    some  silver. 

The  filtrate  of  the  7  per  cent,  solution  showed  a  faint  coloring 
by  sodium  sulphide.  When  using  an  8  per  cent,  solution  the 
coloring  was  still  very  faint.  Molten  silver  chloride  poured 
into  a  5  per  cent,  sodium  chloride  solution  at  100  deg.  F.  decrep- 
itated into  fine  powder;  but  the  filtrate  did  not  show  any  reac- 
tion for  silver. 

The  result  of  this  experiment  shows  that,  if  it  were  possible 
to  regulate  the  base-metal  leaching  in  tanks  so  that  the  outflow- 
ing solution  at  no  time  contains  more  than  6  per  cent,  of  metal 
chlorides,  the  same  would  not  dissolve  any  silver  and  could  be 


TROUGH  LIXIVIATION  223 

allowed  to  run  to  waste,  instead  of  being  subjected  to  a  special 
treatment  to  regain  the  dissolved  silver. 

It  is  different  in  trough  lixiviation.  There  we  have  it  in  our 
power  to  regulate  the  grade  of  concentration  of  the  resulting 
base-metal  solution  at  will,  and  we  can  therefore  produce  at  once 
a  solution  which  is  sufficiently  dilute  not  to  dissolve  silver  chloride. 
This  is  one  of  the  great  advantages  of  trough  lixiviation.  Should 
the  ore  contain  sufficient  cupric  chloride  to  make  the  saving  of 
the  copper  an  object,  the  whole  resulting  solution  may  be 
passed  through  a  series  of  tanks  or  deep  troughs  filled  with  scrap 
iron.  Its  diluted  state  will  not  prevent  the  chemical  reaction. 

I  have  made  a  very  interesting  observation  in  respect  to  the 
solubility  of  the  base-metal  chlorides  with  regard  to  the  trough 
principle.  Two  samples  of  complex  ores  from  different  mining 
districts  were  roasted  with  8  per  cent,  of  salt.  Twenty  grams  of 
each  were  placed  on  a  paper  filter  and  leached  with  water.  When 
all  the  soluble  base-metal  chlorides  had  been  extracted,  the  fil- 
trate was  weighed  in  both  instances,  and  it  showed  that  the 
amount  of  water  required  for  sample  No.  1  was  three  times  the 
weight  of  the  ore,  while  for  sample  No.  2  it  was  seventeen  times 
the  weight  of  the  ore.  By  treating  the  same  quantity  of  ore  one 
and  a  half  minutes  on  the  trough  principle,  no  perfect  extraction 
of  the  base  metals  could  be  obtained,  until  the  same  proportion 
of  water  and  ore  was  used  as  that  required  in  common  leaching, 
viz.:  3  to  1  for  ore  No.  1,  and  17  to  1  for  ore  No.  2.  Both  these 
ores  were  rather  base.  This  behavior  of  the  soluble  chlorides 
toward  the  water  as  solvent  is  undoubtedly  remarkable.  If 
for  both  ores  we  assume  the  filtering  property  to  be  alike,  it  would 
follow,  according  to  the  quantity  of  water  required,  that  if  leached 
in  tanks  ore  No.  2  will  have  to  be  leached  5§  times  as  long  as  ore 
No.  1  to  extract  the  soluble  salts.  The  experiment,  however, 
showed  that  this  extraction  can  be  accomplished  from  both  ores 
in  the  same  short  time  by  bringing  them  at  once  in  contact 
with  their  respectively  required  quantity  of  solvent. 

Care  has  to  be  exercised  in  taking  samples  in  trough  lixivia- 
tion. The  final  residue  sample  can  be  taken  from  the  tank  with 
sampling  irons  in  the  usual  way  before  the  residues  are  dis- 
charged, but  samples  required  for  the  observation  of  the  process 
during  operation  have  to  be  taken  in  a  different  way.  The  whole 
stream  has  to  be  caught  in  a  vessel  of  proper  size,  say  in  an  enam- 


224  HYDROMETALLURGY  OF  SILVER 

eled  iron  kettle  of  two  or  three  gallons  capacity.  The  proper 
place  to  take  the  sample  is  where  the  stream  leaves  the  trough 
and  drops  into  the  settling-tank.  The  kettle  is  quickly  pushed 
under  the  stream  so  that  it  receives  the  whole  stream,  and  quickly 
withdrawn  as  soon  as  it  is  filled.  With  the  proper  mark 
attached  to  it  the  kettle  is  left  undisturbed  until  the  liquid  be- 
comes perfectly  clear,  then  it  is  carefully  decanted  and  filled  up 
again  with  water  and  stirred  well.  This  washing  is  done  twice, 
when,  finally,  the  settled  pulp  is  evaporated  to  dryness,  well  mixed 
and  quartered  down. 

In  the  laboratory,  experiments  can  be  made  on  the  principle 
of  trough  lixiviation  by  introducing  20  grams  of  roasted  ore, 
which,  if  the  test  is  for  silver,  has  previously  been  washed,  into  a 
graduated  cylinder  of  1000  c.c.,  in  which  is  contained  100  c.c. 
or  200  c.c.  of  sodium  hyposulphite  solution,  according  to  the 
proportion  which  is  intended  to  be  used.  The  top  of  the  cylin- 
der has  to  be  tightly  closed  with  the  palm  of  the  hand,  and  the 
cylinder  has  to  be  brought  into  a  horizontal  position,  and  then 
oscillated  in  order  to  make  the  ore  and  solution  pass  quickly  from 
one  end  to  the  other,  to  imitate  the  current  in  a  trough.  This 
is  done  for  about  three-quarters  of  a  minute  or  for  one  minute, 
then  the  contents  of  the  cylinder  are  poured  into  a  filter,  washed 
with  water  to  displace  the  silver  solution  from  the  sand  and  paper, 
dried  and  assayed.  The  same  operations  are  required  in  experi- 
menting with  the  base-metal  chlorides,  except  that  water  is  used 
instead  of  sodium  hyposulphite  solution. 

Having  by  experiment  found  the  required  proportion  of 
water  and  ore  and  solution  and  ore,  the  size  of  pumps,  pipes, 
outlets,  etc.,  can  be  calculated. 

THE  TROUGHS 

A  triangular  shape  of  the  troughs  is  preferable,  because  the 
same  quantity  of  water  will  display  more  energy  for  moving  the 
sand  than  on  a  flat  bottom.  An  inclination  of  three-fourths  of 
an  inch  per  foot  is  sufficient.  We  have  seen  that  when  the  pulp 
passes  through  a  trough  12  to  15  ft.  long  the  extraction  is  ,com- 
pleted,  therefore  no  particular  attention  has  to  be  paid  to  the 
length  of  the  troughs.  The  necessary  trough  connection  from 
feeder  to  sluice-tank  and  from  there  to  the  settling-tank  give 
more  than  enough  length.  The  stream  in  the  trough  moves 


TROUGH  LIXIVIATION  225 

swiftly,  and  therefore  very  little  bottom  pressure  will  be  exerted, 
so  that  it  is  very  easy  to  construct  these  troughs  so  that  they  will 
be  tight. 

SLUICE-TANKS  AND  SLUICING 

As  the  ore  has  to  be  moved  from  the  tanks  by  a  stream  of 
water,  it  is  not  advantageous  to  give  the  tanks  a  large  diameter. 
Twelve  or  14  ft.  is  sufficient,  though  if  circumstances  demand  it 
larger  tanks  can  be  used. 

Figure  52  represents  the  vertical  section,  Fig.  53  the  horizontal 
view,  and  Fig.  54  the  front  view  of  a  settling-tank  arranged  for 
sluicing.  In  the  center  of  the  bottom  is  the  discharge  opening, 
6  in.  in  diameter.  The  cast-iron  discharge-tube,  k,  of  the  same 
inside  diameter,  tightly  fastened  to  the  outside  of  the  tank  bottom, 
corresponds  with  the  discharge-hole.  The  lower  end  of  the  tube 
is  at  right  angles  to  the  upper  end,  and  provided  with  flange  o. 
The  valve  ra,  which  is  provided  with  a  rubber  gasket,  can  be 
pressed  tightly  against  flange  o  by  turning  the  wheel  F.  Flange 
o  and  valve  ra  are  made  of  brass.  Fig.  55  shows  in  detail  the 
construction  of  wheel  F.  Part  of  the  valve-stem  is  square  and 
rests  at  ra  in  a  square  box,  so  that  by  turning  the  wheel  F  the 
valve  m  does  not  turn  too,  but  moves  forward  or  backward.  By 
this  arrangement  the  life  of  the  rubber  gasket  is  much  lengthened, 
as  no  turning  force  is  exercised  against  the  flange  o,  but  only  a 
quiet  pressure.  Around  the  discharge  opening,  and  fastened 
to  the  bottom  of  the  tank,  is  the  wooden  polygon  v,  in  which  is 
cut  the  groove  pr  Around  the  inner  periphery  of  the  tank, 
and  high  enough  to  give  the  filter  bottom  an  inclination  of  at 
least  three-quarters  of  an  inch  to  the  foot,  is  the  groove  p.  Fig.  53 
illustrates  the  construction  of  the  filter  bottom,  which  is  made 
in  sections.  The  filter  cloth  is  well  fastened,  and  kept  in  place 
by  driving  tightly  a  rope  into  the  grooves  pt  and  p.  The  air- 
escape  pipe  d,  which  reaches  to  the  rim  of  the  tank,  enters  the 
latter  close  under  the  filter  bottom.  A  piece  of  hose  is  fastened 
to  the  upper  end  and  can  be  closed  by  a  hose  clamp.  In  Fig.  54 
#!,  q,  are  solution  outlets;  s,  filter  outlet.  Connecting  pipes  g  and 
h  (Figs.  52  and  54)  have,  like  the  discharge-tube  k,  to  be  well 
coated  with  asphaltum  varnish.  In  the  same  figures  z  is  the 
water-pipe;  n,  the  central  hose,  which  ought  to  be  very  stiff  to 
resist  the  pressure  of  the  ore;  it  reaches  down  into  the  discharge- 


226 


HYDROMETALLURGY  OF  SILVER 


SETTLING  VAT  •-   HORIZONTAL  VIEW-  Scale;- Jfct-1  ft. 

FIGS.  52  and  53.  — SETTLING-TANK  ARRANGED  FOR  SLUICING. 


TROUGH   LIXIVIATION 


227 


tube  k,  where  it  has  .to  remain  during  the  process  of  charging. 
Before  charging  the  tank  the  discharge-tube  is  filled  with  water 
through  the  central  hose,  in  order  to  keep  the  latter  filled  with 
water,  which  will  prevent  the  inside  of  the  hose  from  being  ob- 
structed by  ore.  In  the  base-metal  department  the  central  hose 
has  to  be  connected  with  both  the  water-  and  solution-pipe. 
The  connection  with  the  solution-pipe  serves  for  sluicing  the  ore, 


FIG.  54.— SETTLING-TANK,  FRONT  VIEW. 
Scale,  i  in.  =  1  ft. 

while  the  water  connection  is  used  after  the  tank  is  empty  to  free 
the  tank  and  filter  cloth  from  all  adhering  sodium  hyposulphite 
solution  by  rinsing,  otherwise  it  would  get  into  the  base-metal 
solution  and  dissolve  some  silver.  In  the  silver  department 
only  the  connection  with  the  water-pipe  is  required,  as  the 
hose  is  used  only  for  sluicing  out  the  residues. 

When  a  tank  is  ready  to  be  discharged,  the  wheel  F  is  turned, 
and  thus  the  valve  m  pulled  back.     The  water  is  injected  through 


228 


HYDROMETALLURGY  OF  SILVER 


the  central  hose,  while  the  latter  is  gently -moved  up  and  down. 
The  stream  undermines  the  tightly  packed  sand,  causes  a  con- 
tinual caving  in,  until  a  funnel-shaped  opening  is  made  through 
its  depth  to  the  surface.  Then  several  streams  are  made  to  play 
on  the  top,  while  the  central  hose,  with  checked  stream,  is  left 
in  position  to  avoid  obstruction  of  the  discharge-tube  by  a  too 
sudden  rush  of  sand. 

The  central  position  of  the  discharge  opening  and  the  funnel 
shape  of  the  filter  permit  a  quick  and  clean  sluicing.  The  pulp 
leaving  the  discharge-tube  enters  the  sluice-trough  t  underneath 
the  tank,  which  leads  to  the  tailings-trough  u  in  the  silver  depart- 


FIG.  55.  — WHEEL  FOR  CLOSING  DISCHARGE  GATE. 
Scale,  i  in.  =  1  in. 

ment,  or  to  the  silver-leach  trough  in  the  base-metal  department. 
The  charge  being  sluiced  out,  tank  and  filter  have  to  be  well 
rinsed;  the  valve  of  branch  pipe  a  is  opened,  and  the  adhering 
sand  washed  off  from  flange  o  and  valve  ra  by  the  double  sprinkler 
w.  During  discharge  the  valve  m  has  to  be  pulled  back  far 
enough  to  prevent  the  outflowing  pulp  from  striking  it,  otherwise 
the  rubber  gasket  would  soon  wear  out.  Then  the  discharge 
valve  is  closed  again,  and  the  tank  is  ready  to  be  connected  with 
the  other  tanks.  In  Figs.  52  and  54  y  represents  the  silver- 


TROUGH  LIXIVIATION  229 

leach  trough;  b,  the  intersecting  box  above  the  tank,  and  i  a  hose 
made  of  duck. 

ARRANGEMENT  AND  OPERATIONS 

The  construction  and  manipulation  of  the  sluicing  tank  was 
treated  previously  to  the  description  of  the  general  arrangement 
and  operations  of  a  trough  lixiviation  plant  in  order  to  make  it 
better  understood. 

Solution  is  performed  outside  the  tanks  in  troughs,  while  the 
ore  is  moving  in  and  with  the  stream  of  solvent,  and  the  tanks 
are  used  only  to  separate  the  solids  from  the  liquid.  The  system 
is  a  continuous  one;  but  as  the  lixiviation  process  requires  two 
solvents,  first,  water  for  the  removal  of  the  base-metal  chlorides, 
and  then  a  solution  of  sodium  hyposulphite  for  the  extraction  of 
silver,  it  has  to  be  divided  into  two  departments,  the  base-metal 
and  the  silver  departments.  Fig.  56  shows  a  complete  arrange- 
ment. The  upper  series  of  tanks  represents  the  base-metal,  the 
lower  series  the  silver  department.  The  tanks  in  each  depart- 
ment are  placed  on  the  same  level  and  close  together.  They  are 
connected  by  pipes  a,  6,  c,  d,  e,  and  /,  in  such  a  way  as  to  form  a 
perfect  circuit.  These  connecting  pipes  are  placed  a  few  inches 
below  the  rim  of  the  tanks  and  also  on  a  level.  The  diameter  of 
these  pipes  depends  on  the  daily  capacity  of  the  works  and  the 
proportion  of  solvent  and  ore  to  be  used.  Each  tank  has  two 
outlet  pipes  qv  q,  Fig.  54,  on  a  level  with  the  communicating  pipes, 
and  one  S,  from  under  the  filter.  They  all  discharge  into  the 
base-metal  solution  trough  which  leads  outside  the  building, 
either  to  waste  or  to  scrap-iron  tanks  for  the  recovery  of  copper. 
Each  of  these  tanks  is  arranged  and  constructed  for  sluicing,  as 
described  above. 

The  silver-leaching  department  consists  of  the  same  number 
of  tanks  of  the  same  size  and  construction  as  those  of  the  base- 
metal  department,  but  they  are  placed  on  a  lower  level.  They  are 
also  connected  with  communicating  pipes  forming  a  circuit. 

On  a  still  lower  level  are  placed  the  precipitation  tanks,  which 
in  this  case  are  agitated  by  mechanical  stirrers,  and  on  the  next 
platform  we  find  the  filters  for  the  precipitate.  Below  this  are 
the  solution-pump  vats. 

The  triangular  trough  extends  above,  and  in  the  center 
line  of  each  row,  and  wherever  the  trough  branches  off,  there  is 


230 


HYDROMETALLURGY  OF  SILVER 


FIG.  56.  — SYSTEM  FOR  CONTINUOUS  TROUGH   LIXIVIATION. 
For  a  working  capacity  of  50  tons  per  day. 


TROUGH  LIXIVIATION  231 

inserted  an  intersecting  box  which,  if  located  above  a  tank,  is 
provided  with  a  plug-hole  in  the  bottom,  through  which  the  pulp 
flows  if  that  tank  is  to  be  filled.  The  troughs  branching  off  from 
the  intersecting  boxes  are  each  provided  with  a  sliding  gate, 
which  is  kept  closed  except  in  that  trough  into  which  the  stream 
is  to  be  directed. 

The  operations  are  as  follows:  The  roasted  ore  from  the  cool- 
ing-floor is  charged  into  an  iron  hopper  which  at  its  bottom  has 
an  adjustable  worm  discharge.  The  worm  discharges  into  a 
stream  of  water  flowing  into  a  triangular  trough  which  feeds  the 
pulp  into  a  grinding  machine,  in  which  the  lumps  are  mashed, 
and  which  will  be  explained  further  on.  Passing  this  lump- 
mashing  machine,  the  pulp  enters  the  triangular  base-metal 
leach-trough,  and  is  thus  conveyed  automatically  to  the  base- 
metal  tanks.  The  first  tank  to  be  filled  is  tank  C,  Fig.  56;  but 
before  commencing  operation,  the  following  preparations  have 
to  be  made.  The  slide-gate  of  the  trough  leading  from  the 
intersecting  box  to  tank  C  is  opened,  while  those  of  the  other 
two  troughs  are  closed.  The  connecting  pipe  c,  between  tanks  C 
and  G,  is  closed  by  a  plug  from  the  inside  of  C.  The  outlets  of 
all  the  tanks  are  closed  except  the  outlets  ql  q,  near  the  rim  of 
tank  G,  Fig.  54.  Likewise  all  outlets  from  under  the  filters  are 
closed.  Then  the  central  hose  n,  Fig.  52,  is  inserted  into  the  dis- 
charge-tube, the  valve  of  which  is  also  closed.  This  done,  the 
pulp  is  permitted  to  flow  into  tank  C,  Fig.  56.  When  the  solution 
reaches  the  level  of  the  connecting  pipes,  it  will  flow  through 
connecting  pipe  b  into  tank  B,  and  when  B  is  filled  into  tank  A, 
and  so  on  until  the  last  tank  G  is  filled,  when  the  solution  will 
leave  the  tank  through  the  two  outlets  near  the  rim.  As  soon 
as  tank  C  is  filled  with  ore,  the  pulp  is  made  to  flow  into  tank  B. 
The  connecting  pipe  6  is  closed,  and  thus  tank  C  is  disconnected 
from  the  circuit.  The  outlet  s,  below  the  filter  of  tank  C,  is 
opened,  and  the  solution  is  allowed  to  drain  into  the  base-metal 
solution  trough.  When  the  solution  begins  to  disappear  below 
the  surface  of  the  ore,  water  is  admitted,  to  press  out  the  solution 
absorbed  by  the  ore.  When  this  is  done,  sodium  hyposulphite 
solution  is  applied  to  press  out  the  water.  As  soon  as  the  liquid 
flowing  out  through  filter  outlet  s  shows  indications  of  silver,  the 
outlet  s  is  closed,  and  the  charge  is  ready  to  be  sluiced  for  silver 
leaching.  While  tank-charge  C  is  under  the  described  treatment, 


232  HYDROMETALLURGY  OF  SILVER 

which  does  not  take  much  time,  tank  B  is  gradually  filling  with 
ore.  When  filled,  the  pulp  is  made  to  enter  tank  A,  and  tank  B 
is  disconnected  from  the  circuit,  and  its  charge  is  treated  in  the 
same  way  as  that  of  C. 

When  the  charge  C  has  been  prepared  as  described  above, 
it  is  sluiced  out  with  sodium  hyposulphite  solution  to  extract  the 
silver.  Underneath  the  tank  the  pulp  enters  the  silver  leach- 
trough  and  flows  down  to  tank  C,  of  the  silver  department.  The 
arrangement  and  construction  of  the  tanks  are  the  same  as  in 
the  base-metal  department,  and  when  operations  have  been 
started  the  connections  have  to  be  set  the  same  as  described  for 
base-metal  leaching. 

Charge  C,  being  sluiced  out  with  solution,  the  filter-outlet  s 
is  turned  into  the  silver  leach-trough,  the  hose  clamp  is  opened, 
and  the  solution  contained  under  the  filter  is  allowed  to  flow 
out,  and  then  the  inside  of  the  tank  and  the  filter  is  rinsed  with 
water,  which  is  also  allowed  to  flow  into  the  silver  leach-trough. 
This  accomplished,  outlet  hose  s  is  closed  and  returned  to  the 
base-metal  solution  trough.  Then  the  plug  of  connecting  pipe 
c  is  removed,  and  tank  C  is  restored  again  to  the  circuit.  C  being 
empty,  the  flow  in  the  base-metal  solution  trough  will  cease  until 
the  tank  is  filled  again  with  base-metal  solution. 

By  using  the  proper  proportion  of  solvent  and  ore,  the  latter 
will  drop  as  tailings  into  the  tank,  while  all  the  silver  chloride 
which  can  be  extracted  by  the  common  chlorination  test  will  be 
dissolved  by  the  solution.  It  is  well  to  have  a  special  solution 
pipe  by  which  a  stream  can  be  made  to  flow  directly  into  the 
silver  leach-trough  close  to  the  tank,  so  that  if  the  volume  of 
solution  used  in  sluicing  should  not  be  sufficient,  the  proportion 
can  be  maintained  by  the  use  of  this  extra  pipe. 

The  clear  solution  leaving  the  last  tank  Gf  flows  through  the 
silver  solution  trough  into  the  distributing-box,  and  from  there 
to  any  desired  precipitating  tank.  The  bottom  of  the  distribut- 
ing-box has  twice  as  many  holes  as  there  are  precipitating 
tanks.  In  these  holes  bent  lead  pipes  are  inserted  from  below 
and  are  fastened  by  flanges.  Stiff  two-inch  hose-pipes,  of  which 
two  lead  to  each  precipitation  tank,  are  attached  to  these  pipes. 
The  holes  can  be  closed  by  long  plugs.  This  arrangement  I 
found  quite  convenient.  The  operator  can  direct  the  stream 
from  the  main  working  floor,  without  being  obliged  to  creep  over 


TROUGH  LIXIVIATION  233 

all  the  precipitating  tanks,  as  is  customary  in  many  lixiviating 
works. 

When  one  tank  is  filled  with  tailings,  it  is  disconnected  from 
the  circuit  and  the  pulp  is  admitted  into  the  next  tank.  Outlet 
sv  under  the  filter,  is  opened;  the  solution  still  contained  in  the 
tank  is  allowed  to  drain  off,  and  the  part  retained  by  absorption 
is  displaced  by  water.  The  tailings  are  then  sluiced  with  water. 
Where  water  is  scarce,  the  wash-water  can  be  collected  and  used 
for  sluicing  out  the  final  tailings. 

It  is  of  advantage  to  connect  the  solution-pipe  with  the  solu- 
tion-pump. In  the  first  place  a  higher  pressure  can  be  obtained 
than  from  the  storage  tanks,  and  on  the  other  hand  by  the  num- 
ber of  pump-strokes  per  minute  the  volume  of  solution  can  be 
calculated.  The  extra  stream  into  the  trough  outside  the  tank, 
if  one  should  be  used,  it  is  better  to  have  from  the  storage  tank,  as 
the  high  pressure  from  the  pump  would  splash  in  the  trough. 

PRECIPITATING  VATS 

Figs.  57  to  61  illustrate  the  construction  of  precipitating 
vats  and  a  convenient  arrangement  of  filters  of  the  precipitate 
which  I  devised  and  have  operated.  This  construction  of  tank 
and  filters  is  particularly  adaptable  for  works  which  are  not 
equipped  with  air-compressor  and  filter-presses.  The  tanks  are 
provided  with  machine-stirrers,  of  the  construction  indicated 
in  the  drawings.  The  stirrer  s  (Figs.  57  and  58)  has  to  make 
about  30  r.p.m.  if  the  diameter  of  the  tank  is  not  more  than 
8  or  9  ft.  It  is  set  in  motion  or  stopped  by  working  the  friction- 
clutch  /  (Fig.  57).  The  wings,  g  (Fig.  58),  which  reach  near  to 
the  bottom,  are  about  3  in.  wide  and  are  kept  in  position  by 
triangular  pieces  of  boards.  They  break  the  violent  current 
around  the  periphery  and  throw  the  solution  toward  the  center, 
thus  causing  a  strong  whirling  motion.  In  Figs.  58  and  59,  a  is 
the  discharge-pipe,  or  decanter,  for  the  clear  solution,  and  b  for 
the  precipitate.  Pipe  d,  Fig.  59,  passes  in  front  of  all  the  precipi- 
tation tanks,  and  conveys  the  calcium  or  sodium  sulphide  solu- 
tion from  the  reservoir  to  the  tanks.  The  branch  pipe  c  reaches 
above  the  rim  of  the  tank  and  ends  in  a  rubber  hose,  which  is 
provided  with  a  clamp.  In  precipitating,  the  stream  can  be 
conveniently  regulated  by  the  use  of  the  clamp,  and  the  operator, 
by  observing  the  color  produced  by  the  precipitant  in  the  moving 


234 


HYDROMETALLURGY  OF  SILVER 


solution,  can  finish  this  operation  in  a  very  short  time,  and  much 
more  easily  than  by  using  buckets. 

One  man  can  precipitate  three  tanks  at  a  time  without  assist- 
ance. The  solution  is  so  thoroughly  agitated  that  a  very  perfect 
separation  of  the  silver  sulphide  takes  place.  The  separation 
is  so  perfect  that  the  bottom  of  the  tank  can  distinctly  be 


FIG.  57.— PRECIPITATION  TANK,  VERTICAL  SECTION. 

seen  through  5  ft.  of  solution.  To  produce  a  quick  and  perfect 
separation  of  the  precipitated  silver  sulphide,  the  solution  has 
to  be  vigorously  agitated.  This  cannot  be  well  done  in  tanks  of 
14  or  16  ft.  in  diameter.  It  is  much  preferable  to  have  smaller 
tanks  and  a  larger  number  of  them.  A  good  size  is  8  to  9  ft.  in 


TROUGH   LIXIVIATION 


235 


diameter  and  6  ft.  deep.  In  some  leaching  works  where  there 
are  large  precipitation  tanks  in  use,  we  find  the  bad  practice  of 
discharging  the  precipitate  only  once  a  week,  in  some  even  only 
once  a  month.  Fresh  precipitate  forms  large  flakes,  which  settle 
easily  and  cleanly.  After  two  or  three  days  it  assumes  a  dry  and 
sandy  condition,  and  if  stirred  up  divides  into  very  fine  particles, 
which  are  kept  suspended  in  the  solution  for  a  long  time.  The 
result  of  such  a  practice  is  that  the  decanted  solution  will  not  be 
free  from  precipitate  when  used  again  for  extraction,  and  a  black 


~~Scale  1A^  I  ft. 
FIG.  58.  — PRECIPITATION  TANK,  PLAN. 

coating  of  precipitate  will  cover  the  surface  of  the  ore  in  the 
leaching  tanks.  The  precipitate  ought  therefore  to  be  removed 
from  the  precipitating  tanks  every  day,  and  should  never  be 
allowed  to  remain  longer  than  two  days.  By  using  smaller 
tanks  with  the  machine-stirrer,  and  by  discharging  the  precipitate 
every  day,  or  every  other  day,  the  circulating  solution  can  be 
kept  so  clear  that,  even  after  a  prolonged  leaching  of  five  or  six 
days,  no  black  coating  can  be  observed  on  the  top  of  the  ore. 
The  machine-stirrer  does  good  work  also  in  discharging  the 


236 


HYDROMETALLURGY  OF  SILVER 


precipitate.  After  the  clear  solution  has  been  decanted  within 
a  few  inches  of  the  precipitate,  the  stirrer  is  set  in  motion,  and  the 
discharge-pipe  b  opened.  The  stirrer  agitates  the  precipitate 


MACHINE 

FOR   HOFMANN'8 

CONTINUOUS  TROUGH  LIXIVIATION 
FRONT  VIEW 


FIG.  59.  — PRECIPITATING  VAT. 

until  nearly  all  is  discharged  into  the  niters,  or  into  the  pressure 
tank  by  which  the  filter-press  is  filled.  Thus  the  cleaning  of  a 
tank  can  be  done  in  a  very  short  time. 


TROUGH  LIXIVIATION 


237 


Where  the  filter-press  is  not  used,  a  filter  arrangement  as  shown 
in  Figs.  59,  60  and  61  will  be  found  very  convenient.  Two  rows 
of  filters  are  so  arranged  that  they  are  in  communication  with 
each  other  by  depressions  cut  into  the  divides  of  the  frame.  By 
allowing  the  precipitate  to  flow  into  one  filter,  all  the  adjoining 
filters  will  gradually  be  filled,  one  after  the  other,  without  requir- 
ing the  attention  of  the  operator.  The  filters,  which  are  made  of 
common  cotton  sheeting,  are  shallow,  and  the  precipitate  will 
thus  be  spread  in  a  comparatively  thin  layer  over  a  large  filter- 
ing surface.  Under  each  row  of  filters  is  placed  a  trough,  n, 
which  receives  the  filtrate  and  conveys  it  to  the  pump-tank  P, 
below  the  floor.  These  filter  frames  are  placed  in  front  of  the 
precipitating  tanks  and  can  be  made  to  contain  quite  a  number 


FIGS.  60  and  61.— FILTER  FRAME. 
Fig.  60  is  a  horizontal  view.    Fig.  61  is  a  section  on  line  A-B. 

of  filters,  as  shown  in  Fig.  56,  where  two  sections  serve  for  five 
precipitating  tanks.  Owing  to  the  shallowness  of  the  filters 
and  the  large  filtering  surface,  the  solution  will  drain  off  fast, 
and  in  five  or  six  hours  the  precipitate  will  be  stiff  enough  to  be 
charged  with  wooden  hand-paddles  into  the  drying  furnace. 

The  extraction  in  trough  lixiviation  is  not  produced  by  fil- 
tration; and  the  silver  as  well  as  the  base-metal  chlorides  con- 
tained in  the  lumps  of  the  roasted  ore  will  not  be  extracted  if 
the  lumps  are  allowed  to  enter  the  trough  together  with  the  finer 
portion.  Actual  grinding  is  not  necessary,  but  in  order  to  ob- 
tain good  results  it  is  necessary  first  to  mash  the  lumps.  An 
agitator  will  not  perform  the  work  rapidly  enough,  and  some 
quick-grinding  machine  has  to  be  employed  for  this  purpose. 


238 


HYDROMETALLURGY  OF  SILVER 


Figs.  62,  63,  64  and  65  represent  a  grinding  machine  specially 
designed  by  myself  for  use  in  trough  lixiviation.  The  mantle  m 
is  concave-shaped  and  is  stationary.  The  muller  /,  a  flatter  cone 
than  the  concave  of  the  mantle,  is  inserted  from  below,  and  can 
be  lowered  or  raised  by  the  screw  g,  for  the  purpose  of  regulating 
the  fineness  to  which  the  lumps  are  intended  to  be  mashed. 


FIG.  62.  — LUMP-GRINDING  MACHINE,  ELEVATION. 

Mantle  and  muller  are  provided  with  exchangeable  shoes  and  dies. 
The  centrifugal  force  developed  by  the  rotation  of  the  muller 
greatly  assists  the  discharge  of  the  pulp.  Toward  the  center 
the  shoes  and  dies  are  provided  with  teeth,  while  toward  the 
periphery  their  surface  is  smooth  (Figs.  64  and  65).  The  teeth 
cut  the  larger  lumps,  the  smaller  ones  are  mashed  by  the  smooth 
part  of  the  cones.  Water  and  ore  are  charged  through  A,  while 


TROUGH  LIXIVIATION 


239 


the  circular  cast-iron  trough  t  receives  the  pulp  and  conveys  it  to 
the  base-metal  leach-trough  o.  The  canvas  strip  k  prevents  the 
pulp  from  flying  over  the  rim  of  the  trough.  By  far  the  main 
portion  of  the  roasted  ore  is  fine  enough  to  pass  through  without 


^i  ft. 


FIG.  63.  — LUMP-GRINDING  MACHINE,  PLAN. 

being  affected,  and  only  a  small  part  will  have  to  be  mashed. 
For  this  reason,  and  on  account  of  the  softness  of  the  wet  material, 
large  quantities  of  ore  can  be  put  through  in  twenty-four  hours 
with  but  very  little  wear  of  the  machine,  which  in  itself  is  of 
simple  and  cheap  construction. 


240  HYDROMETALLURGY  OF  SILVER 


PRACTICE  OF  TROUGH   LIXIVIATION  AT  CUSIHUIRIACHIC. 

The  Don  Enrique  Mining  Company  at  Cusihuiriachic  had 
accumulated  a  large  dump  of  second-class  ore,  which  contained 
about  25  oz.  silver  per  ton,  but  filtered  so  badly  on  account  of  the 
large  amount  of  porphyry  in  the  gangue  that  it  was  not  profitable 
to  work  it  by  tank  lixiviation.  After  the  large  lixiviation  works 
of  this  company  were  destroyed  by  fire,  I  undertook  to  work  this 
second-class  ore  by  trough  lixiviation,  and  treated  12,000  tons 
by  this  method,  at  a  good  profit. 

The  work  was  done  in  the  old  North  Mexican  mill  buildings, 
which  were  formerly  used  for  tank  lixiviation  until  the  ore  in  the 
company's  mine  gave  out.  There  were  two  sluicing-vats  with 


-i  ft.  ^--_r  %-i  ft. 

FIG.  64.  —  Mantle.  FIG.  65.  —  Muller. 

LUMP-GRINDING  MACHINE. 

central  discharge  and  flat,  funnel-shaped  filter  bottoms,  which 
were  placed  about  10  ft.  above  the  rim  of  jthe  old  leaching-vats,  of 
which  there  were  eight,  arranged  in  two  rows.  Each  vat  measured 
14  ft.  in  diameter  and  3J  ft.  in  depth.  These  tanks  were  used  as 
settling-tanks,  and  were  connected  each  with  the  other  by  a 
4-in.  pipe  inserted  near  the  rim,  thus  forming  a  circuit  of  all  the 
vats,  so  that  if  one  tank  was  filled  the  solution  could  flow  into 
the  other,  and  from  there  into  the  next,  and  so  on.  The  filter 
bottoms  of  these  vats  were  covered  with  sheeting  and  a  4-in. 
layer  of  washed  river  sand.  On  the  cooling-floor  a  small  hopper 
with  funnel-shaped  bottom  was  erected  and  covered  with  an 
inclined  screen  of  J-in.  mesh,  to  prevent  the  lumps  from  entering. 
The  hopper  discharged  to  a  short  screw  conveyor,  and  the  con- 
veyor into  the  cups  of  a  belt  elevator.  The  elevator  lifted  the 


TROUGH  LIXIVIATION  241 

roasted  ore  and  discharged  it  into  a  short  triangular  trough,  in 
which  a  stream  of  water  was  running.  The  trough  was  so 
arranged  that  the  pulp  could  be  conveyed  to  either  of  the  two 
sluicing-tanks. 

The  speed  of  the  screw  conveyor  being  always  the  same,  the 
feed  of  roasted  ore  into  the  trough  was  uniform,  and  the  desired 
proportion  of  ore  and  water  was  easily  regulated  by  the  stream 
of  water.  When  the  first  vat  was  filled  the  pulp  was  made  to 
enter  the  second  sluicing-vat.  While  the  second  vat  filled,  the 
liquid  in  the  first  vat  became  clear  and  was  drawn  off  by  siphons, 
so  that  when  vat  No.  2  was  full  vat  No.  1  was  ready  to  receive 
again  the  stream  of  pulp.  This  was  repeated  until  both  vats 
were  fairly  filled  with  washed  ore.  After  the  first  charge  of  pulp 
the  outlet  under  the  filter  of  each  vat  was  opened,  discharging  a 
clear  stream  while  filling  was  in  progress,  the  ore,  which  had  by 
this  time  settled  on  the  bottom,  acting  as  filter.  This  increased 
the  filling  capacity  of  the  vats.  When  both  tanks  were  charged 
the  solution  above  the  ore  was  allowed  to  drain,  and  then  water 
was  applied  to  displace  the  solution  absorbed  by  the  ore. 

The  silver-leaching  troughs  were  not  longer  than  was  necessary 
to  reach  to  each  of  the  eight  settling-vats.  The  intersecting 
boxes  over  each  settling-vat  were  18  in.  square  and  12  in.  deep, 
with  a  3-in.  round  opening  in  the  bottom,  which  could  be  closed 
with  a  wooden  plug.  To  prevent  splashing  during  charging  a 
4-in.  canvas  hose  was  fastened  around  the  opening  in  the  bottom, 
reaching  down  to  within  a  few  inches  below  the  rim  of  the  settling- 
vat. 

Shortly  before  sluicing  began  the  plug  in  the  box  above  the 
first  tank  was  removed  and  a  piece  of  filter  cloth,  kept  in  place 
by  several  bricks,  was  spread  directly  under  the  canvas  hose,  to 
protect  the  sand  filter.  The  pulp,  dropping  always  on  the  same 
place  in  the  vat,  the  ore,  or  rather  the  residues  (because  the  silver 
is  already  extracted  when  the  pulp  drops  into  the  vat),  will  form 
a  cone,  which,  however,  will  never  project  much  above  the  sur- 
face of  the  solution,  because  the  material  in  the  solution,  being 
loose  and  lighter  than  the  material  above  the  solution,  will  slide 
down.  Thus  a  tank  of  14  to  16  ft.  diameter  will  be  charged 
pretty  evenly.  To  fill  the  lower  space  around  the  periphery  a 
short  trough  was  placed  under  the  drop,  and  by  changing  the 
position  of  the  trough  gradually  the  stream  was  directed  to  all 


242  HYDROMETALLURGY  OF  SILVER 

points.  Long  before  the  vat  was  filled  with  residues  the  solution 
reached  the  level  of  the  communicating  pipe  and  flowed  into  the 
next  tank  (No.  2),  and  then  into  No.  3  and  No.  4.  When  tank 
No.  2  was  filled  with  solution  the  outlet  under  the  sand  filter 
was  opened  and  a  clear  stream  of  solution  was  discharged.  The 
same  was  done  when  No.  3  was  filled  with  solution,  and  so  on. 
The  filtration  was  very  free,  and  the  volume  of  solution  entering 
vat  No.  3  was  much  reduced,  and  still  more  so  the  overflow  into 
No.  4,  so  that  the  solution  seldom  occupied  more  than  three  vats 
besides  the  one  which  was  undergoing  charging.  This  gave 
ample  time  to  treat  and  discharge  the  residues,  and  to  renew  the 
filter,  if  necessary,  before  the  vat  was  again  required  in  the  cir- 
cuit. By  providing  each  tank  with  a  sand  filter  only  clear  fil- 
tered solution  entered  the  precipitating  vats,  and  no  overflowing 
solution  at  all. 

When  tank  No.  1  was  filled  sufficiently  with  residues  the 
stream  of  pulp  was  changed  to  flow  into  vat  No.  2  by  opening  the 
corresponding  plug-hole  and  closing  the  one  above  No.  1.  The 
communicating  pipe  between  vat  No.  1  and  No.  2  was  closed, 
and  the  solution  allowed  to  drain  and  then  displaced  by  water. 
The  leaching  with  water  was  continued  until  the  outflowing 
stream  did  not  show  any  reaction  for  silver.  When  a  vat  was 
partly  filled  with  residues  the  outlet  under  the  filter  could  be 
opened,  giving  a  clear  stream,  though  the  vat  was  still  under- 
going the  process  of  filling.  When  vat  No.  1  was  disconnected 
from  the  circuit  and  No.  2  subjected  to  the  operation  of  charg- 
ing, the  overflowing  solution  moved  one  tank  farther  and  com- 
menced to  fill  No.  5;  when,  in  the  course  of  the  operation,  the 
overflowing  solution  reached  vat  No.  8,  vat  No.  1  was  empty  and 
prepared  to  receive  the  overflowing  solution.  To  prevent  the 
overflowing  stream,  when  entering  an  empty  tank,  from  washing 
away  the  filter  sand,  a  piece  of  filter  cloth  was  spread  over  that 
part  of  the  filter  and  tacked  to  the  inside  of  the  vat,  so  that  the 
stream  which  flowed  down  on  the  side  of  the  vat  did  not  do  any 
harm  to  the  filter. 

Samples  of  the  pulp  taken  at  the  drop  showed  that  all  the 
silver  chloride  was  extracted  while  the  pulp  was  flowing  through 
the  trough,  and  that  the  ore  actually  dropped  as  spent  residues 
into  the  vat.  A  similar  experience  was  realized  with  regard  to 
the  base-metal  salts.  After  the  pulp  (ore  and  water)  passed 


TROUGH  LIXIVIATION  243 

through  the  short  trough  from  the  elevator  to  the  sluicing-vat, 
it  was  found  by  samples  taken  at  the  drop  that  all  the  heavy 
metal  salts  soluble  in  water  had  been  dissolved,  and  that  only 
some  sodium  sulphate  still  remained.  Regular  samples  of  the 
base-metal  solution  were  taken,  but  never  found  to  contain  any 
silver. 

TROUGH  LIXIVIATION  EXPERIMENTS  ON  A  LARGE  SCALE 

While  investigating  the  metallurgical  problem  of  the  lead- ' 
zinc  ores  of  the  San  Francisco  del  Oro  ore  near  Parral,  Chihuahua, 
Mexico,  experiments  were  also  made  to  treat  the  roasted  ore  by 
the  trough-lixiviation  method.  These  experiments  were  made 
previous  to  the  working  of  the  second-class  ore  dumps  at  Cusi- 
huiriachic,  described  above.  Experiences  gained  by  this  experi- 
ment were  used  to  advantage  in  working  the  "Cusi"  ore. 

The  Bosque  mill,  near  Parral,  was  very  inconveniently 
arranged.  The  main  inconvenience  was  the  want  of  grade; 
therefore,  the  locality  did  not  permit  the  erection  of  a  complete 
system  for  trough  lixiviation,  and  the  experiments  had  to  be  made 
with  only  one  circuit  of  six  tanks,  and  I  was  obliged  to  use  the 
same  troughs  and  tanks  for  base-metal  and  afterwards  for  silver 
leaching.  The  washed  ore  had  to  be  removed  from  the  tanks 
and  brought  to  the  head  of  the  trough  for  silver  leaching.  Not- 
withstanding this  inconvenience,  the  experiments  gave  very 
interesting  results  and  information. 

By  a  triangular  trough,  138  ft.  in  length,  f-in.  fall  per  foot, 
with  a  feed-box  at  the  upper  end,  and  intersected  by  five  square 
boxes,  the  pulp  could  be  conveyed  to  any  of  the  six  tanks  of  the 
circuit.  The  tanks  were  connected  by  pipes  inserted  near  the  rim. 

The  ore  used  in  this  experiment  was  roasted  in  the  modified 
Howell  furnace.  It  was  charged  into  a  running  stream  of  water 
at  the  rate  of  64  tons  per  twenty-four  hours.  The  pulp  passed 
through  the  whole  length  of  trough  in  fifty-five  seconds. 

In  order  to  find  out  how  much  of  the  base-metal  salts  were 
dissolved  during  this  short  time,  and  to  ascertain  the  required 
length  of  trough,  samples  were  taken  at  different  places,  dried 
and  then  subjected  to  a  thorough  washing  in  the  laboratory, 
with  the  following  results: 

Roasted  ore  before  t roughing  contained  12  per  cent,  in  salts 
soluble  in  water. 


244  HYDROMETALLURGY  OF  SILVER 

(1)  The  sample  taken  after  the  pulp  passed  the  entire  length 
of  138  ft.  still  contained  in  salts  soluble  in  water  4.9  per  cent. 

(2)  The  sample  taken  after  the  pulp  passed  through  58  ft. 
of  trough  still  contained  in  soluble  salts  4.5  per  cent. 

(3)  The  sample  taken  after  the  pulp  passed  through  12  ft. 
of  trough  still  contained  in  soluble  salts  3.6  per  cent. 

The  above  results  are  just  in  reverse  order  from  what  would 
be  expected;  but  it  was  not  possible  to  take  the  sample  from  the 
same  portion  of  moving  pulp,  which  may  account  for  this  irregu- 
larity. Though  this  assumption  is  rather  arbitrary,  we  may 
accept  it  for  want  of  a  better  explanation.  If  we  take  the  aver- 
age of  the  three  results,  we  find  that  the  pulp  after  troughing 
still  contained  4.7  per  cent,  of  salts  soluble  in  water,  or,  as  the 
roasted  ore  before  washing  contained  12  per  cent,  of  such  salts, 
that  60.8  per  cent,  can  be  extracted  while  the  pulp  passes  through 
12  ft.  of  trough,  or  in  4.7  seconds.  Long  troughs  are,  therefore, 
not  essential  for  base-metal  leaching.  In  order  to  ascertain  if, 
in  tank  Iixiviati6n  in  the  usual  routine,  a  larger  percentage  of 
the  soluble  salts  is  extracted,  a  sample  was  taken  from  a  tub 
charge  after  it  had  been  washed  for  eight  hours  and  was 
ready  for  silver  leaching.  The  outflowing  water  gave  with  cal- 
cium sulphide  only  faint  white  clouds,  the  usual  indication  that 
base-metal  leaching  is  completed.  The  sample,  after  drying  and 
weighing,  was  subjected  to  a  second  washing  in  the  assay  office, 
and  the  result  showed  that  of  the  original  percentage  of  soluble 
salts  61.7  per  cent,  were  extracted  by  leaching  in  the  tanks,  which 
is  only  0.9  per  cent,  more  than  was  extracted  in  the  trough  in  4.7 
seconds. 

In  both  cases  about  the  same  percentages  of  soluble  salts  are 
retained  by  the  ore,  which  only  by  a  prolonged  leaching  can  be 
removed.  They  are  not  heavy  metal  salts,  but  principally 
sodium  sulphate  and  sodium  chloride.  In  the  present  case 
mostly  sodium  sulphate;  for  an  analysis  of  the  stock  solution, 
after  three  months'  use,  showed  it  to  contain  only  0.098  per  cent, 
chlorine,  while  the  white  clouds  produced  by  an  addition  of  cal- 
cium sulphide  proved  to  be  gypsum. 

TIME  REQUIRED  FOR  BASE-METAL  LEACHING 

Though  the  dissolving  of  the  base  metals  is  almost  instan- 
taneous, considerable  time  is  consumed  in  preparing  the  charge 


TROUGH  LIXIVIATION  245 

for  silver  leaching,  caused  principally  by  the  time  required  to 
press  out  the  base-metal  solution  by  water;  this  time  was  found 
to  be  3  hours  and  25  minutes  for  a  charge  of  8.39  tons.  However, 
the  total  time  is  still  3  hours  and  35  minutes  less  than  in  tank 
lixiviation. 

The  time  is  divided  as  follows: 

IN  TROUGH  LIXIVIATION 

hrs.  min. 

Leaching  and  filling  the  tanks 3  6 

To  drain  remaining  solution  from  top  of  the  ore -  34 

To  press  out  base-metal  solution  by  water   3  25 

To  press  out  water  by  hypo  solution _! 20 

Total  time 8  h.      25  m. 

IN  TANK  LIXIVIATION 

hrs. 

Charging 3 

Base-metal  leaching 8 

Pressing  out  with  hyposulphite  solution    1 

Total  time 12  hrs. 

QUANTITY  OF  WATER  REQUIRED 

Sufficient  water  had  to  be  used  to  make  the  pulp  move  freely 
through  the  trough,  and  to  produce  a  sufficiently  diluted  base- 
metal  solution  in  order  not  to  dissolve  any  silver  chloride.  The 
results  were  obtained  with  702  gallons  per  ton  of  ore,  which  is 
equivalent  to  about  one  weight  of  ore  to  three  of  water.  When 
the  tank  was  charged,  and  clean  water  turned  on  to  press  out 
the  solution,  the  speed  of  filtration  was  12  inches  per  hour  in  a 
tank  of  10  ft  2  in.  diameter,  which  is  equivalent  to  2065  gallons  in 
3  hours  25  minutes  for  a  charge  of  8.39  tons,  or  246  gallons  per 
ton.  After  silver  leaching,  it  took  2  hours  30  minutes  to  press 
out  the  hyposulphite  solution.  Summing  up,  we  find  the  total 
consumption  of  water  as  follows : 

gal. 

In  troughing 702 

Pressing  out  the  base-metal  solution  by  water  .   246 
Pressing  out  the  hyposulphite  solution  by  water    181 

Total  consumption  per  ton 1129  =  150.5  cu.  ft. 

In  tank  lixiviation  the  consumption  of  water  was  found  to  be 
703  gallons  per  ton  of  ore,  or  93.7  cu.  ft.,  which  shows  an  increased 
consumption  in  trough  lixiviation  of  56.8  cu.  ft.  per  ton. 


246  HYDROMETALLURGY  OF  SILVER 


QUANTITY  OF  SILVER  DISSOLVED  BY  THE  BASE-METAL  SOLUTION 

One  liter  of  the  702  gallons  of  base-metal  solution  was  precipi- 
tated with  calcium  sulphide.  The  precipitate,  after  fluxing  and 
treating  like  a  common  ore  assay,  returned  not  more  than  0.0002 
grams  fine  silver.  If  one  liter  contains  0.0002  grams  silver,  702 
gallons  will  contain  0.532  grams,  which  is  the  total  amount  of 
silver  dissolved  from  the  whole  charge  of  8.39  tons  of  ore,  or  0.06 
grams,  equal  to  0.002  oz.  silver  per  ton.  This  is  practically 
nothing,  and  the  wash-water  can  therefore  be  allowed  to  run  to 
waste,  without  causing  any  perceptible  loss  in  silver. 

SILVER  LEACHING 

After  three  tanks  were  filled,  and  the  base-metal  solution 
pressed  out  with  water,  and  the  water  with  a  0.38  per  cent,  sodium 
hyposulphite  solution,  the  ore  was  allowed  to  drain.  Then  the 
ore  was  shoveled  out  and  removed  for  silver  leaching  to  the  head 
of  the  trough.  Being  at  a  time  rather  late  in  the  evening,  the 
ore,  saturated  with  hyposulphite  solution,  was  left  in  a  pile  over 
night.  The  next  morning,  however,  it  was  found  that  some  of 
the  silver  chloride  was  decomposed  during  the  night,  and  that  the 
chlorination  test  tailings  had  increased  from  5.24  oz.  to  9.03  oz. 
per  ton.  Taking  the  tailings  value  of  9.03  oz.  per  ton  as  basis, 
the  experiment  was  continued.  The  measured  stream  of  solu- 
tion was  kept  uniform,  while  the  rapidity  of  charging  the  ore 
was  changed  according  to  the  desired  proportion.  In  order  to 
ascertain  the  proper  length  of  trough,  samples  were  taken  at 
different  places,  with  the  following  results: 

Assay  office  chlorination  tailings  of  the  pulp,  9.03  oz.  per  ton; 
strength  of  solution,  0.38  per  cent.;  proportion,  one  weight  of  ore 
to  five  of  solution;  rate  of  working,  38  tons  of  ore  per  day. 

(1)  Sample  taken  from  spout  of  feed-box  when  entering  the 
trough:  Tailings,  9.8  oz.  per  ton. 

(2)  Sample  taken  after  passing  12  ft.  of  trough:  Tailings,  8.13 
oz.  per  ton. 

(3)  Sample  taken  after  passing  70  ft.   of  trough:  Tailings, 
8.85  oz.  per  ton. 

(4)  Sample  taken  after  passing  100  ft.  of  trough:  Tailings, 
8.13  oz.  per  ton. 


TROUGH  LIXIVIATION  247 

(5)  Sample  taken  after  passing  120  ft.  of  trough:  Tailings, 
8.60  oz.  per  ton. 

(6)  Sample  taken  after  passing  138  ft.  of  trough,  while  drop- 
ping in  tank:  Tailings,  8.60  oz.  per  ton. 

This  experiment  gave  the  very  surprising  information  that 
actually  only  a  few  feet  of  trough  are  required  to  produce  a  per- 
fect dissolving  of  the  silver  chloride.  It  shows  that  the  passing 
through  12  ft.  of  trough,  or  in  4.7  seconds,  the  extraction  is  com- 
plete, and  that  longer  troughs  are,  therefore,  not  necessary. 
This  is  of  importance,  as  it  simplifies  the  construction  of  trough- 
lixiviating  works  and  reduces  the  required  grade. 

OTHER  PROPORTIONS,  WORKING  THE  SAME  LOT  OF  ORE 

(7)  Proportion;  1  ore  to  3.4  solution;  working  rate,  55.8  tons 
per  twenty-four  hours;  tailings,  7.89  oz.  per  ton. 

(8)  Proportion,  1  ore  to  2|  solution;  working  rate,  84.5  tons 
per  twenty-four  hours;  tailings,  9.56  oz.  per  ton. 

(9)  Proportion,  1  ore  to  10  solution;  working  rate,  19.05  tons 
per  twenty-four  hours;  tailings,  9.09  oz.  per  ton. 

These  results  show  that  the  proportion  of  1  ore  to  3.4  solution 
gave  the  best  results,  the  tailings  being  1.14  oz.  per  ton  poorer 
than  the  chlorination  assay  called  for. 

A  second  series  of  experiments  was  made,  and  particular 
attention  was  paid  to  avoid  the  decomposition  of  silver  chloride. 
The  charges  were  subjected  to  silver  leaching  soon  after  being 
saturated  with  hyposulphite  solution. 

Chlorination  test  tailings,  5.25  oz.  per  ton. 

Strength  of  solution,  0.50  per  cent. 

(1)  Proportion,  1   ore  to  3.4  solution;  tailings,  3.59  oz.   per 
ton. 

(2)  Proportion,  1  ore  to  6  solution;  tailings,  3.8  oz.  per  ton. 
These  are  very  satisfactory  results;  the  tailings  are  as  poor  as, 

in  fact  poorer  than,  those  obtained  in  tank  lixiviation  after  four 
days'  silver  leaching.  The  proportion  1  to  3.4  proved  again  to 
be  sufficient,  producing  tailings  1.66  oz.  poorer  than  the  chlorina- 
tion test  called  for;  the  quantity  of  solution  required  for  this  ore 
is,  therefore,  very  moderate,  much  less  than  that  required  in 
tank  lixiviation. 


248  HYDROMETALLURGY   OF  SILVER 

QUANTITY  OF  SOLUTION  REQUIRED 

By  using  the  proportion  1  : 3.4  we  need  100.8  cu.  ft.,  or  816 
gal.  of  solution  to  circulate  for  each  ton  of  ore.  In  tank  lixivia- 
tion  the  required  quantity  for  this  ore  was  found  to  be  658  cu.  ft., 
or  4935  gal.  for  each  ton  of  ore,  or  about  six  times  as  much  as  in 
trough  lixiviation. 

TIME  REQUIRED  FOR  SILVER  LEACHING 

hrs.       min. 

Troughing  and  filling  the  tank    3          36 

Draining  the  solution  from  the  top  of  the  ore    -  34 

Pressing  out  the  solution  with  water    ^2 30 

Total  time 6  h.      4m. 

In  the  following  we  will  compare  the  total  time  required  by 
the  two  methods  from  the  time  the  ore  enters  the  leaching  works 
until  it  is  ready  for  discharge. 

TIME  REQUIRED  IN  TANK  LIXIVIATION 

hrs. 

Charging   2 

Base-metal  leaching 8 

Pressing  out  the  water  by  solution 1 

Silver  leaching   96 

Pressing  out  the  solution  by  water li 

Total  time lOSYhours. 

TIME  REQUIRED  IN  TROUGH  LIXIVIATION 

hrs.       min. 

Base-metal  leaching  and  filling  the  tank 3  6 

To  drain  the  wash-water  from  top  of  ore   —  34 

To  press  out  the  base-metal  solution  by  water 3  25 

To  press  out  the  water  by  hyposulphite  solution  ...    1  20 

Silver  leaching  (sluicing  with  solution)    3  36 

Draining  solution  from  top  of  ore —  34 

Pressing  out  the  solution  by  water    _2 30 

Total  time 15  h.        5m. 

To  work  a  charge  of  Del  Oro  ore  by  the  trough  system  takes 
15  hours  5  minutes,  while  by  tank  lixiviation  it  takes  108  hours 
30  minutes,  or  about  seven  times  as  long. 

FINENESS  OF  THE  PRECIPITATE 

The  burned  silver  precipitate  contained  20.9  per  cent,  of  fine 
silver,  while  that  obtained  in  tank  lixiviation  during  the  same 
week  and  from  the  same  lot  of  ore  contained  only  17  per  cent, 
fine  silver. 


TROUGH  LIXIVIATION  249 


ADVANTAGES  OF  TROUGH  LIXIVIATION 

It  can  clearly  be  seen  that  trough  lixiviation  offers  many 
advantages  over  tank  lixiviation,  especially  in  large  works,  or  if 
badly  filtering  or  slowly  extracting  ore  has  to  be  treated.  In 
large  works,  where  twenty  or  more  leaching- vats  are  in  operation, 
each  one  in. a  different  stage  of  the  process,  much  attention  is 
required  to  avoid  mistakes,  while  in  trough  lixiviation  care  has 
to  be  given  to  only  a  few  tanks.  If  the  ore  filters  badly  it  will 
take  a  very  long  time  to  extract  the  silver  by  leaching  in  tanks, 
while  in  troughs  the  silver  as  well  as  the  base-metal  salts  dissolve 
almost  instantaneously,  and  the  effect  of  the  bad  filtering  will 
be  felt  only  while  the  charge  is  draining  and  the  solution  is 
being  displaced  by  water.  It  makes  the  treatment  of  very  badly 
filtering  ore  possible,  which  otherwise  could  not  be  treated  by 
lixiviation. 

The  most  important  advantage  of  trough  lixiviation  is  the 
fact  that  this  method  enables  the  operator  to  bring  the  ore  into 
sudden  contact  with  any  desired  quantity  of  the  solvent.  This 
is  a  very  important  fact,  as  it  offers  the  means  to  do  away  with 
the  special  treatment  of  the  base-metal  solution,  because  a  solu- 
tion of  this  can  be  made  sufficiently  diluted  not  to  dissolve 
any  silver  chloride.  Furthermore,  the  extraction  of  silver  from 
lead-bearing  ores  is  slow,  and  requires  an  extensive  plant.  It  is 
well  known  to  leachers  of  such  ores  that,  while  the  main  portion 
of  the  silver  is  extracted  in  a  short  time,  the  remaining  few  ounces 
will  be  tenaciously  retained  by  the  ore.  Thus  it  happens  that, 
while  the  main  portion  of  the  silver  can  be  extracted  in  the  first 
six  or  eight  hours,  the  remaining  eight  or  ten  ounces  of  extract- 
able  silver  will  require  three  or  four  days,  sometimes  more.  I 
have  seen  cases  in  which  only  one  ounce  per  ton  in  every  twenty- 
four  hours  of  prolonged  lixiviation  could  be  extracted.  This 
very  singular  phenomenon  is  difficult  to  explain.  It  seems  that 
only  that  part  of  the  silver  is  so  difficult  to  extract  which  origi- 
nally was  contained  in  the  lead-mineral  of  the  ore.  I  have  observed 
that  whenever  the  galena  of  the  ore  became  richer  in  silver,  or 
had  increased  in  quantity,  the  extraction  became  slow  and  drag- 
ging. This,  together  with  the  fact  that  the  main  portion  of  the 
silver  can  be  quickly  extracted,  indicates  that  the  slowness  of 
the  extraction  is  not  principally  due  to  the  disadvantageous  in- 


250  HYDROMETALLURGY  OF  SILVER 

fluence  of  lead  sulphate  on  the  dissolving  energy  of  the  solution 
for  silver  chloride,  but  that  it  must  be  due  to  some  other  cause, 
which  prevents  the  silver  contained  in  the  lead  ore  from  dissolving 
quickly  except  in  large  volumes  of  sodium  hyposulphite  solution. 
In  troughs,  when  the  ore  is  brought  at  once  into  contact  with  the 
required  volume  of  solution,  the  silver  dissolves  almost  instan- 
taneously, and  the  presence  of  lead  ore  does  not  retard  trough 
lixiviation:  it  merely  entails  the  use  of  larger  quantities  of  solu- 
tion, which  is  undoubtedly  a  very  advantageous  feature  of  this 
method. 

The  cost  of  a  plant  is  much  less  if  arranged  for  troughs,  as 
large  quantities  of  ore  can  be  treated  in  a  comparatively  small 
plant. 


XVIII 

THE  RUSSELL  AND  KISS  PROCESSES 

(1)  THE  RUSSELL  PROCESS 

THE  Russell  process,  which  was  *so  elaborately  and  well  written 
up,  and  about  which  so  many  statements  of  excellent  results 
were  published,  and  which  was  in  consequence  thereof  introduced 
at  several  places  with  a  large  expenditure  of  capital,  has  not 
proved  a  success.  As  this  process  attracted  much  attention  and 
has  found  its  way  into  all  metallurgical  text-  and  hand-books,  it 
is  interesting  and  instructive  to  investigate  the  cause  of  the 
failure. 

It  was  claimed  that  the  "  extra  solution  "  —  a  solution  of  a 
double  salt  of  cuprous  hyposulphite  and  sodium  hyposulphite, 
manufactured  by  adding  a  solution  of  copper  sulphate  to  a 
solution  of  sodium  hyposulphite  —  exerted  a  highly  energetic 
dissolving  and  decomposing  action  upon  metallic  silver,  silver 
sulphide,  silver  minerals  belonging  to  the  group  of  antimonial  and 
arsenical  sulphides,  and  other  silver  combinations.  Based  on  this 
property  of  the  "  extra  solution, "  it  was  claimed  that  silver  ores 
treated  by  this  process  required  a  less  careful  chloridizing  roast- 
ing, or  only  an  oxidizing  roasting,  and  that  even  raw  sulphureted 
ores  could  be  successfully  desilverized. 

These  claims  attracted  general  attention,  and  if  they  had 
been  true  in  the  sense  in  which  they  were  given  out  this  process 
would  have  marked  a  decided  step  forward  in  the  hydrometallurgy 
of  silver;  but  these  claims  were  based  on  results  obtained  on  a 
very  small  scale,  and  under  conditions  which  are  not  practicable 
to  create  and  maintain  on  a  large  scale,  while  the  results,  even 
under  such  conditions,  especially  with  regard  to  raw  ores,  proved 
to  be  not  good  enough  to  justify  the  application  of  the  process  in 
practical  metallurgy.  To  protect  himself  against  failure  it  is  of 
the  greatest  importance  that  the  experimenter  should  execute 

251 


252  HYDROMETALLURGY  OF  SILVER 

his  experiments  under  conditions  which  can  be  practically  main- 
tained on  a  large  scale.  As  soon  as  he  has,  by  working  on  a  large 
scale,  to  change  the  conditions  wrhich  he  maintained  in  his  labora- 
tory tests,  he  will  obtain  very  different  results,  and  only  too 
often  will  experience  a  complete  failure. 

If  a  half  ounce  of  raw  sulphureted  ore  is  treated  in  a  beaker 
with  a  large  excess  of  a  32  per  cent,  "extra  solution"  for  twelve 
hours  with  frequent  stirring,  as  was  done  by  Russell  in  the  labora- 
tory, it  is  experiment  ing  under  condit  ions  which  cannot  be  profitably 
maintained  on  a  large  scale,  and  the  results  thus  obtained,  if  pub- 
lished, should  be  given  as  a  matter  of  scientific  interest  but  not  as 
actual  results  of  a  new  metallurgical  process,  especially  if  the  con- 
dition under  which  these  results  were  obtained  are  withheld.  More 
or  less  silver  will  be  dissolved  by  such  a  strong  "extra  solution"; 
a  32  per  cent,  sodium  hyposulphite  solution  applied  in  the  same 
manner  will  also  dissolve  some  silver;  but  it  is  not  practicable  to 
work  with  such  concentrated  solutions,  since  they  cannot  be  main- 
tained without  a  very  large  consumption  of  sodium  hyposulphite 
and  copper  sulphate.  Besides,  raw  sulphureted  ores  do  not  filter 
well,  and  the  ore  would  have  to  be  treated  in  agitating  tanks  for 
twelve  hours,  and  the  separation  of  the  solution  from  the  residues 
would  have  to  be  effected  by  means  of  large  filter-presses.  These 
manipulations,  together  with  the  very  large  consumption  of 
chemicals,  would  cost  more  than  roasting  with  salt.  The  weakest 
part  of  this  method  as  applied  to  raw  ore,  however,  is  that  the 
extraction  at  the  best  is  so  inferior  and  incomplete  that  its  appli- 
cation is  entirely  out  of  the  question. 

More  favorable  results  can  be  obtained  by  treating  raw  oxi- 
dized ores,  or  sulphureted  ores  which  were  first  subjected  to  a 
thorough  oxidizing  roasting,  wherein  there  are  not  so  serious  dif- 
ficulties preventing  its  application  on  a  large  scale,  though  only 
in  exceptional  cases  will  it  be  rational  to  employ  it.  A  dilute  so- 
lution can  be  used,  and  often  50  to  70  per  cent,  of  the  silver  will 
be  extracted.  But  similar  results  can  be  obtained  by  using  a 
straight  solution  of  sodium  hyposulphite.  Some  of  the  San 
Francisco  del  Oro  ore,  a  highly  sulphureted  lead-zinc  ore,  was 
roasted  oxidizingly  in  a  reverberatory  furnace  by  me,  and  part 
of  it  treated  with  sodium  hyposulphite  and  part  with  Russell's 
"extra  solution."  The  roasted  ore  contained  29.3  oz.  silver  per 
ton.  By  leaching  with  sodium  hyposulphite  17.2  oz.  silver  were 


THE  RUSSELL  AND  KISS  PROCESSES  253 

extracted,  while  with  Russell's  "extra  solution"  the  extraction 
gave  17.64  oz.  or  0.43  oz.  silver  per  ton  more.  But  an  increased 
extraction  of  less  than  half  an  ounce  of  silver  per  ton  does  not 
justify  the  extra  consumption  of  7  Ib.  of  copper  sulphate  and  5 
Ib.  of  sodium  hyposulphite,  which  amount  is  stated  to  be  the 
consumption  of  chemicals  in  the  Russell  process  per  ton  of  ore 
treated.  It  would  be  folly  to  oxidize  an  ore  for  lixiviation  and 
obtain  an  extraction  of  only  50  to  70  per  cent,  instead  of  chlori- 
dizing  it  and  obtaining  an  extraction  of  90  per  cent,  and  more, 
especially  since  by  the  modern  method  of  chloridizing  roasting 
the  loss  of  silver  by  volatilization  is  greatly  reduced,  and  does 
not  exceed  the  loss  occurring  in  oxidizing  roasting.  The  claim 
that  ores  require  only  an  oxidizing  roasting,  if  treated  by  the 
Russell  process,  is  therefore  a  mistake,  and  cannot  be  verified  by 
actual  and  satisfactory  working  results. 

It  was  mentioned  above  that  only  in  exceptional  cases  can 
lixiviation  of  raw  ores  be  executed  successfully.  One  of  the 
cases  is  when  in  oxidized  ore  the  silver  occurs  as  chloride.  How- 
ever, it  is  more  advantageous  to  subject  such  ore  first  to  a  short 
red  heat  to  melt  the  silver  chloride.  Chloride  of  silver,  as  it 
occurs  in  nature,  is  very  dense  and  dissolves  very  slowly,  but  if 
the  ore  is  heated  the  silver  chloride  will  melt  and  impregnate  the 
surrounding  ore  particles,  in  which  condition  it  offers  a  large  sur- 
face to  the  solvent  and  permits  a  much  quicker  extraction.  Oxi- 
dized ores  in  which  the  silver  occurs  as  antimonate  can  also  be 
successfully  leached.  Low-grade  oxidized  ores,  which  will  yield 
60  to  70  per  cent,  by  leaching  raw,  can  also  be  successfully  treated 
if,  on  account  of  the  small  tenor  of  silver  in  the  ore,  the  increased 
amount  extracted  by  chloridizing  does  not  exceed  the  cost  of 
roasting.  However,  the  use  of  sodium  hyposulphite  for  such 
ores  will  be  found  more  economical  than  the  use  of  the  "extra 
solution,"  the  additional  cost  of  chemicals  not  being  covered  by 
the  slight  gain  in  extraction. 

After  the  process  was  tried  on  a  large  scale  without  success, 
Russell  modified  his  process.  He  adopted  chloridizing  roasting 
and  leaching  with  straight  sodium  hyposulphite,  and  after  this 
was  done  applied  his  "extra  solution,"  by  which  he  claimed  to 
obtain  a  much  better  extraction.  But  this  assertion  was  again 
based  on  laboratory  experiments,  which  were  made  under  dif- 
ferent conditions  than  could  be  maintained  in  the  works.  Based 


254  HYDROMETALLURGY  OF  SILVER 

on  these  results  it  was  next  claimed  that  silver  ores  require  a 
less  careful  chloridizing  roasting,  because  the  "extra  solution" 
dissolves  the  unchloridized  part  of  the  silver.  But  this  claim 
was  also  found  to  be  an  illusion,  as  has  been  demonstrated  by 
many  failures  on  a  large  scale.  The  following  case  is  an  illustra- 
tion: 

In  Sombrerete,  Zacatecas,  Mexico,  the  lixiviation  process 
with  sodium  hyposulphite  was  in  successful  operation  for  years, 
until  a  new  company  was  organized  to  work  the  property  on  a 
larger  scale.  A  new  and  large  mill  was  erected  to  suit  the  re- 
quirements of  the  Russell  process.  The  success  in  the  old  mill 
was  based  on  a  good  chloridizing  roasting  in  reverberatory  fur- 
naces. The  results,  however,  were  entirely  different  when  the 
new  mill  was  set  in  operation.  In  this  the  ore  was  roasted  in 
a  Stetefeldt  furnace,  which  is  not  suitable  for  such  heavy  sul- 
phureted  ore,  and  the  chlorination  was  far  from  being  satisfac- 
tory. This  gave  an  opportunity  to  demonstrate  the  claim  that 
the  "extra  solution"  exerts  an  energetic  dissolving  and  decom- 
posing action  on  the  unchloridized  part  of  the  silver.  However, 
the  "extra  solution"  failed  to  react,  and  after  eleven  months  of 
unsuccessful  trials  the  company  failed  and  the  property  changed 
hands. 

The  new  company  abandoned  the  Stetefeldt  furnace  and  the 
Russell  process,  built  a  suitable  number  of  reverberatory  fur- 
naces, and  adopted  the  common  lixiviation  process  with  sodium 
hyposulphite  with  great  success.  The  mill  is  still  in  operation 
and  treats  regularly  60  to  80  tons  of  ore  per  day. 

The  Cusi  company  at  Cusihuiriachic,  Chihuahua,  Mexico, 
had  a  similar  experience.  The  old  lixiviation  process  was  for 
years  in  successful  operation  on  a  scale  of  50  to  60  tons  per  day. 
Induced  by  the  glowing  representations  of  the  advantages  of 
the  Russell  process,  the  company  adopted  it,  but  after  1J  years' 
trial,  with  heavy  financial  loss,  the  process  was  discarded  and 
lixiviation  with  sodium  hyposulphite  was  resumed. 

(2)  THE  Kiss  PROCESS 

This  process  was  recommended  for  the  mutual  extraction  of 
silver  and  gold  from  sulphureted  auriferous  silver  ores.  Kiss 
subjected  the  ore  to  a  chloridizing  roasting,  leached  with  water 
to  remove  the  base-metal  chlorides,  and  extracted  the  silver  and 


THE  RUSSELL  AND  KISS  PROCESSES  255 

gold  with  a  solution  of  calcium  hyposulphite.  As  precipitant 
he  used  calcium  polysulphide.  According  to  Kiss,  in  chloridizing 
roasting  subchloride  of  gold  is  formed,  which  is  neither  soluble  in 
water  nor  in  a  solution  of  sodium  chloride,  but  is  soluble  in  a  solu- 
tion of  calcium  hyposulphite.  This  method  was  in  actual  opera- 
tion in  several  places  in  Hungary,  and  while  the  extraction  of 
the  silver  was  satisfactory  (90  per  cent.)  the  extraction  of  the 
gold  varied  greatly  at  different  places;  from  90  to  20  per  cent. 
This  was  undoubtedly  caused  by  the  way  the  roasting  was  done 
at  the  different  places.  Subchloride  of  gold  cannot  exist  at  a 
temperature  prevailing  in  the  furnace,  especially  not  if  the  tem- 
perature toward  the  end  is  increased,  as  was  formerly  generally 
done;  it  is  then  decomposed  into  metallic  gold  and  chlorine. 
The  subchloride  of  gold,  however,  is  formed  if  the  ore  after  being 
discharged  from  the  furnace  is  allowed  to  cool  slowly.  If  the 
hot  ore  on  leaving  the  furnace  is  dumped  into  a  pile,  or  better  in 
a  bin,  the  generation  of  chlorine  continues  down  to  a  tempera- 
ture which  is  low  enough  not  to  decompose  the  subchloride  of 
gold,  and  therefore  opportunity  is  given  for  the  formation  of 
this  gold  combination.  (This  has  been  discussed  in  another 
chapter  of  this  treatise.)  The  variation  of  the  gold  extraction 
at  the  different  places  was  therefore  most  likely  due  more  to 
whether  the  chloridized  ore  was  cooled  slowly  or  quickly  rather 
than  to  the  different  character  of  the  ores. 

We  have  seen,  too,  that  calcium  hyposulphite  in  contact  with 
sodium  sulphate  forms  sodium  hyposulphite  and  calcium  sul- 
phate (gypsum),  which  precipitates,  so  that  Kiss  actually  did  not 
work  with  calcium  hyposulphite  after  the  solution  was  used  for 
some  time,  but  with  sodium  hyposulphite,  as  the  lime  salt  was 
decomposed  by  coming  in  contact  with  the  remaining  sodium 
sulphate  of  the  roasted  ore.  It  is  questionable  whether  calcium 
hyposulphite  is  a  more  energetic  solvent  for  subchloride  of  gold 
than  sodium  hyposulphite;  at  least,  it  was  not  proved  by  Kiss' 
method,  as  he  actually  leached  with  the  soda  and  not  with  the 
lime  salt. 


XIX 

THE  AUGUSTIN  PROCESS 

IN  this  process  the  material  is  roasted  with  salt.  The 
resulting  silver  chloride  is  dissolved  by  a  hot  concentrated  brine. 
From  the  solution  the  silver  is  precipitated  by  copper,  and  the 
copper  by  iron.  The  remaining  solution  is  freed  from  iron  and 
sodium  sulphate  and  brought  into  circulation  again. 

The  ore  is  roasted,  first  oxidizingly  to  transform  the  metal  sul- 
phides into  oxides,  and  the  silver  sulphide  as  much  as  possible 
into  silver  sulphate,  so  that  during  the  chloridizing  period  but 
little  volatile  or  soluble  metal  chlorides  will  be  formed.  The 
material  which  conforms  best  with  these  requirements  is  copper 
matte.  The  oxidizing  roasting  is  continued  until  the  iron  has 
changed  into  red  oxide  and  the  copper  into  cupric  oxide.  If  no, 
or  not  enough,  sulphates  are  left  at  the  end  of  the  oxidizing  period 
to  decompose  the  salt,  ferrous  sulphate  has  to  be  added 
with  the  salt  to  generate  the  chlorine.  The  roasted  material  is 
not  first  leached  with  water,  but  the  hot  concentrated  salt  solu- 
tion is  applied  at  once  to  the  ore.  If  not  all  the  copper  was 
changed  into  oxide,  and  some  chloride  or  subchloride  is  still  pres- 
ent, this  will  be  dissolved  by  the  salt  solution,  and  the  subchlo- 
ride oxidizes  during  the  precipitation  of  the  silver  to  basic  insoluble 
chloride  of  copper  and  makes  the  precipitated  silver  impure. 
If  the  matte  contains  lead  the  hot  brine  will  dissolve  lead  chloride, 
which  in  cooling  precipitates  as  white  flakes  and  also  makes  the 
silver  impure.  Such  material  has  to  be  leached  first  with  hot 
water. 

The  stream  of  hot  brine  coming  from  the  leaching-vats  is 
divided  in  many  small  streams,  each  of  which  is  made  to  strike 
metallic  copper,  which  is  placed  in  small  wooden  tubs  provided 
with  a  filter  and  an  outlet  below  the  filter.  Under  these  tubs  is 
placed  another  row  of  similar  tubs,  and  below  this  still  another, 

256 


THE  AUGUSTIN  PROCESS  257 

so  that  the  brine  passes  through  quite  a  number  of  such  tubs 
filled  with  copper.  They  are  arranged  on  benches.  Leaving 
the  last  row,  the  desilverized  solution  is  conveyed  to  a  series  of 
tanks  for  the  precipitation  of  the  copper  by  scrap  iron. 

An  advantageous  feature  of  this  process  is  that  the  silver  is 
precipitated  in  the  metallic  state;  from  a  solution  of  sodium  hypo- 
sulphite, the  silver  could  not  be  precipitated  with  copper  without 
a  decomposition  of  the  sodium  hyposulphite.  This  method, 
however,  has  many  disadvantages  which  make  it  unfit  for  work 
on  a  really  large  scale.  The  dissolving  energy  for  silver  chloride 
of  a  concentrated  hot  salt  solution  is  much  inferior  to  that  of 
sodium  hyposulphite;  it  takes  68  parts  of  sodium  chloride  to  dis- 
solve one  part  of  silver  chloride,  while  it  takes  only  two  parts  of 
sodium  hyposulphite.  The  handling  of  large  quantities  of  hot 
concentrated  brine  is  rather  unclean  and  troublesome.  All  parts 
of  the  vats  and  tubs  above  the  solutions  and  the  whole  of  the 
outside  become  covered  with  incrustations  of  salt;  besides,  it  is 
exceedingly  difficult  to  keep  them  tight,  as  such  a  concentrated 
and  hot  solution  finds  its  way  even  through  the  fibers  of  the  wood. 

Cement  copper  acts  more  energetically  in  the  precipitation  of 
the  silver  than  scrap  copper,  but  the  resulting  cement  silver  is 
not  quite  as  clean,  as  it  usually  contains  more  copper  particles, 
which  have  to  be  removed  by  treating  with  hydrochloric  acid. 


XX 


EXTRACTION  WITH  SULPHURIC  ACID 

THIS  process,  like  the  Augustin  method,  is  exclusively  used 
for  the  extraction  of  the  silver  from  products  of  smelting,  prin- 
cipally black  copper  and  copper  matte. 

(1)    EXTRACTION  OF  SILVER  FROM  COPPER  MATTE 

The  products  of  this  process  are  cupric  sulphate  (blue  vitriol, 
blue  stone),  which  goes  in  solution,  and  residues  in  which  the 
silver  and  gold  remain,  from  which  they  are  finally  extracted  by 
smelting  with  lead  ores. 

(a)  The  Old  Method 

Besides  the  extraction  of  the  precious  metals,  it  is  of  great 
importance  that  the  resulting  cupric  sulphate  should  be  as  free 
from  iron  as  possible,  because  the  blue  vitriol  is  bought  in  the 
market  as  such,  and  if  it  contains  ferrous  sulphate  its  value  is 
much  reduced.  To  accomplish  this,  the  common  copper  matte 
has  to  be  first  concentrated  by  repeated  roasting  and  smelting, 
not  only  to  enrich  it  in  copper  but  to  free  it  as  much  as  possible 
from  iron.  Though  red  oxide  of  iron  dissolves  much  more 
slowly  in  diluted  sulphuric  acid  than  cupric  oxide,  the  resulting 
copper  solution  will  contain  by  far  too  much  ferrous  sulphate  to 
produce  a  marketable  blue  vitriol  if  the  copper  matte  contains 
a  large  percentage  of  iron. 

At  Freiberg,  Saxony,  where  this  process  is  in  operation,  they 
concentrate  the  original  copper  matte  of  38  to  44  percent,  copper 
by  repeated  roasting  and  smelting  until  it  contains  70  or  more 
per  cent,  copper  and  only  0.3  per  cent.  iron.  This  matte  is  crushed 
coarse  (J-in.  mesh)  and  roasted  twelve  hours  in  a  reverberatory 
furnace.  The  charge  of  1000  Ib.  is  continuously  stirred,  then 

258 


EXTRACTION  WITH  SULPHURIC  ACID  259 

withdrawn  from  the  furnace,  finely  pulverized  and  subjected 
during  four  hours  to  a  "dead  roasting."  The  roasted  matte  is 
sifted,  the  coarse  pulverized  and  mixed  with  the  fine. 

For  the  treatment  of  the  roasted  matte  with  diluted  sulphuric 
acid  there  are  cylindrical  upright  drums  2  ft.  8  in.  in  diameter 
and  3J  ft.  high,  made  of  i-in.  lead  supported  by  an  iron  frame- 
work, whch  is  covered  with  thin  sheet  lead.  The  drum  has  two 
outlets,  a  lower  one  for  discharging  the  residues  and  one  higher 
up  for  the  discharge  of  the  copper  solution.  Extending  to  near  the 
bottom  there  is  a  lead  pipe  for  the  injection  of  superheated 
steam.  This  pipe  is  so  arranged  that  it  can  be  lowered  and 
raised.  The  drum  is  charged  with  3  cu.  ft.  of  mother  liquor, 
and  3  cu.  ft.  of  common  sulphuric  acid  of  45  to  47  deg.  B., 
which  gives  a  solution  34  to  36  deg.  Beaume".  Superheated  steam 
is  now  applied,  and  when  the  solution  is  boiling,  200  Ib.  of  roasted 
matte  is  gradually  charged  with  a  hand  shovel  through  a  movable 
funnel.  The  pulp  has  to  be  charged  continuously.  After  a 
quarter  of  an  hour,  9  cu.  ft.  of  mother  liquor  is  added  to  the 
contents  of  the  drum,  which  fills  it  to  about  7  in.  below  the  rim. 
The  boiling  with  steam  is  continued  for  4  to  5  hours,  during  which 
time  stirring  has  to  be  done  at  intervals.  The  steam  is  turned  off, 
and  half  an  hour  is  given  for  the  residues  to  settle.  Then  the 
solution  is  drawn  off  through  the  upper  outlet  and  conveyed  to 
settling  tubs,  while  the  residue,  containing  all  the  silver  and 
about  5  per  cent,  copper,  is  transferred  into  a  special  basin,  where 
it  is  washed  with  water.  The  still  hot  copper  solution  is  left 
about  an  hour  in  the  settling-tanks,  during  which  time  some  basic 
iron  salts  settle.  Then  it  is  drawn  off  into  the  crystallizing  tanks. 
The  residues  are  dried  and  transferred  to  the  lead  smelting. 

The  blue  vitriol  crystals  after  five  to  seven  days  are  removed 
from  the  crystallizing  tanks.  They  are  not  of  a  clear  blue  color, 
but  have  a  greenish  appearance  from  a  certain  percentage  -of  iron 
which  they  contain,  and  are  again  dissolved  and  recrystallized. 
All  the  solutions  are  very  acid. 

The  yearly  production  of  blue  vitriol  resulting  from  this  process 
is  about  800  tons. 

(b)  0.  Hofmann's  Method 

We  have  seen  that,  to  treat  copper  matte  with  sulphuric 
acid  by  the  old  method,  it  is  required  to  first  eliminate  nearly 


260  HYDROMETALLURGY   OF  SILVER 

all  the  iron  from  it  by  repeated  roasting  and  smelting,  which  at 
the  same  time  produces  a  matte  very  rich  in  copper.  This 
careful  elimination  of  the  iron  is  done  to  permit  the  production 
of  a  purer  blue  vitriol.  We  have  also  seen  that  in  order  to  avoid 
a  concentration  of  the  resulting  cupric  sulphate  solution  for 
crystallization  the  proportion  of  acid  and  roasted  matte  is  so  regu- 
lated as  to  produce  at  once  a  solution  strong  enough  for  that 
purpose.  If  this  is  done  for  economical  reasons,  to  avoid  the 
cost  of  evaporation,  the  object  is  not  attained,  as  it  necessitates 
recrystallization  because  the  crystals  contain  too  much  iron,  not- 
withstanding the  fact  that  the  matte  by  previous  treatment  con- 
tained but  very  little  iron.  The  solubility  of  red  oxide  of  iron  in 
sulphuric  acid  increases  with  the  strength  of  the  acid,  therefore 
it  is  not  advisable  to  bring  the  matte  in  sudden  contact  with  too 
strong  an  acid.  Cupric  oxide  dissolves  so  much  easier  than  the 
iron  oxide  that  a  weak  acid  can  be  used  which  will  readily  dis- 
solve the  cupric  oxide,  but  very  little  of  the  iron  oxide,  and  the 
resulting  cupric  sulphate  solution  will  be  much  purer.  The 
strong  first  solution  makes  this  method  not  suitable  to  be  worked 
on  a  really  large  scale,  because  concentrated  solutions  require 
but  little  cooling  to  form  fine  crystals,  which  are  very  annoying  in 
the  separating  of  the  residues  from  the  solution,  because  these 
crystals  will  form  in  the  filter-press,  clog  the  filter  cloth  and 
prevent  a  free  filtration.  They  will  form  in  the  mass  of  the 
residues,  whence  they  can  be  removed  only  by  a  prolonged  wash- 
ing with  hot  water,  causing  the  formation  and  accumulation  of 
too  large  quantities  of  a  very  weak  solution. 

I  discovered  a  method  of  purifying  a  cupric  sulphate 
solution  from  iron,  antimony,  arsenic,  etc.,  which,  together  with 
a  modification  of  the  manner  in  which  the  process  of  dissolving 
the  cupric  oxide  is  executed,  made  the  sulphuric  acid  process  not 
only  much  cheaper,  but  also  applicable  for  working  common 
copper  matte  without  previous  concentration,  at  the  same  time 
producing  a  very  pure  product  of  blue  vitriol. 

The  method  is  based  on  the  reaction  that,  if  through  a  hot 
neutral  solution  of  cupric  sulphate  which  contains  ferrous  sulphate 
a  stream  of  air  is  forced,  and  at  the  same  time  cupric  oxide  is  added, 
ferric  oxide  is  precipitated  and  cupric  sulphate  will  go  in  solution. 
Instead  of  specially  prepared  cupric  oxide,  roasted  copper  matte 
is  used,  which  contains  sufficient  cupric  oxide  for  the  reaction. 


EXTRACTION   WITH   SULPHURIC  ACID  261 

The  manner  in  which  the  dissolving  of  the  roasted  matte  is 
performed  is  modified,  inasmuch  as  the  roasted  matte  is  not  added 
to  the  bulk  of  acid  of  sufficient  strength  to  produce  at  once  a 
cupric  sulphate  solution  of  the  required  concentration,  but  the 
solution  is  gradually  enriched  in  copper  until  the  desired  degree 
is  reached.  This  is  done  by  preparing  first  a  3  per  cent,  acid 
solution.  Such  a  weak  acid  dissolves  but  very  little  iron  oxide. 
Then,  when  hot,  the  copper  is  added  gradually  as  a  stream  of 
roasted  matte,  while  at  the  same  time  a  small  stream  of  60  deg. 
B.  acid  flows  into  the  tank.  The  solution  in  the  tank  is  agitated 
vigorously  by  a  machine-stirrer.  The  proportion  of  matte  and 
acid  has  to  be  so  regulated  that  the  solution  always  maintains 
its  strength  of  3  per  cent,  in  acid,  while  the  content  of  copper 
increases.  This  is  controlled  by  frequent  volumetric  tests. 
When  the  desired  strength  in  copper  is  nearly  attained,  the  influx 
of  acid  is  stopped  and  only  so  much  matte  added  as  will  neutralize 
the  3  per  cent.  acid.  Thus  the  copper  matte  does  not  come  in 
contact  with  a  stronger  acid,  and  the  resulting  solution  does  not 
contain  more  than  0.7  to  1  per  cent,  iron,  notwithstanding  that 
the  treated  material  is  very  rich  in  iron  oxide. 

This  process  was  introduced  by  me  in  the  large  smelting 
works  of  the  Consolidated  Kansas  City  Smelting  and  Refining 
Company,  at  Argentine,  Kansas,  for  the  treatment  of  the 
argentiferous  leady  copper  matte,  first  on  a  medjum  large  scale, 
but  on  account  of  its  successful  working  it  underwent  an  enlarge- 
ment every  year,  until  a  daily  producing  capacity  of  60  tons  of 
blue  vitriol  was  reached.  In  the  following  a  short  description 
of  the  process  and  the  apparatus  will  be  given.1 

The  material  treated  is  a  leady  copper  matte  containing  34  to 
40  per  cent,  copper  and  12  to  14  per  cent.  lead.  It  is  crushed  first 
in  a  rock  breaker  and  then  pulverized  in  a  Krupp  ball-mill  of  100 
tons  daily  capacity  through  a  screen  of  50  meshes  to  the  lineal  inch. 
The  roasting  is  done  in  three  two-story  Pearce  turret  furnaces, 
each  provided  with  three  fireplaces  at  the  lower  and  two  at  the 
upper  hearth.  For  the  benefit  of  the  subsequent  operation  the 
roasting  has  to  be  conducted  with  great  care,  and  with  attention 

1  A  more  elaborate  description  can  be  found  in  "  The  Mineral  Industry," 
Vol. 'VIII  and  Vol.  X.  The  plant  is  not  now  in  operation,  the  smelting 
works  with  which  it  was  connected  having  been  closed  down  and  dismantled 
in  1902. 


262  HYDROMETALLURGY   OF  SILVER 

equally  divided  between*  the  oxidation  of  the  copper  and  of  the 
iron.  The  copper  is  to  be  converted  into  cupric  oxide  and  sul- 
phate, and  the  iron  into  red  oxide,  in  which  state  it  dissolves  only 
slowly  in  hot  diluted  sulphuric  acid.  The  formation  of  cupric 
sulphate  is  very  desirable,  as  it  saves  acid  in  the  subsequent 
treatment,  but  still  it  is  not  advisable  to  conduct  the  roasting  so 
that  as  much  as  possible  sulphate  is  formed,  because  the  roasted 
material  will  then  contain  too  much  soluble  iron,  which  would 
make  the  resulting  cupric  sulphate  solution  too  impure.  At  a 
comparatively  early  stage  of  the  roasting  nearly  all  the  copper 
is  in  a  state  in  which  it  can  be  extracted  by  diluted  sulphuric 
acid;  about  75  per  cent,  of  it  is  present  as  oxide  and  25  per  cent,  as 
sulphate.  The  roasting,  however,  cannot  be  considered  complete 
at  this  stage,  because  the  roasted  material  contains  yet  too  much 
soluble  iron. 

At  the  beginning  of  the  operation  the  temperature  should  not 
be  raised  above  that  produced  by  the  combustion  of  the  sulphur 
of  the  matte.  During  this  period  the  charge  assumes  a  rather 
bright  red  appearance,  an  effect  due  more  to  light  than  to  heat, 
and  if  excess  of  air  is  admitted  to  the  furnace  but  few  lumps 
will  form,  even  with  very  leady  copper  matte.  Gradually,  as 
the  oxidation  advances,  the  surface  of  the  charge  becomes  darker, 
especially  near  the  air-doors,  and  when  the  entire  surface  of  the 
charge  begins  to  darken,  the  fire  is  slightly  increased  to  prevent 
cooling,  as  from  this  time  on  the  supply  of  heat  furnished  by  the 
oxidation  decreases  rapidly.  If  this  condition  is  overlooked 
and  the  charge  cools  too  much,  it  will  be  difficult  to  raise  the 
temperature  again  to  the  proper  degree. 

The  roasting  is  continued  at  a  moderate  temperature  until 
no  more  heat  is  evolved  by  the  oxidation  of  the  material,  after 
which  the  temperature  must  be  raised  to  a  cherry  red.  A  skilled 
roaster  can  readily  determine  this  point  by  stirring  the  charge; 
if  the  particles  thrown  to  the  surface  brighten,  the  roasting  has  not 
advanced  far  enough,  but  if  the  entire  charge  presents  a  dead  and 
uniform  red  color,  it  shows  that  this  part  of  the  roasting  has  been 
completed,  and  that  it  is  time  for  an  increase  of  temperature. 
This  can  now  be  done  without  danger  of  lumping  the  charge, 
because  by  this  time  the  greater  part  of  the  sulphides  has  been 
oxidized.  The  increase  of  the  temperature  is  necessary  for  two 
reasons :  first,  to  hasten  the  oxidation  of  the  remaining  sulphides, 


EXTRACTION  WITH  SULPHURIC  ACID  263 

which  would  require  a  very  long  time  at  a  low  temperature,  and 
second,  in  order  to  decompose  the  iron  salts  and  to  convert  them 
into  the  red  oxide.  The  task  for  the  roaster  is  to  convert  as  much 
iron  as  possible  into  the  red  oxide  without  decomposing  the 
cupric  sulphate  present.  Cupric  sulphate  resists  considerable 
heat,  more  so  than  the  ferrous  salts,  and  it  is  possible  to  conduct 
the  roasting  in  this  way;  but  the  increase  of  temperature  requires 
judicious  care,  because  if  the  heat  is  too  high  the  cupric  sulphate 
will  be  reduced  to  cuprous  oxide,  in  which  condition  but  half  of 
the  copper  is  soluble  in  diluted  sulphuric  acid.  If  crystals  of 
cupric  sulphate  are  exposed  to  heat  and  air,  it  will  be  noticed 
that  after  the  acid  is  expelled  the  mass  assumes  a  red  color 
showing  the  formation  of  cuprous  oxide.  If  heating  is  continued, 
it  turns  black  by  being  oxidized  to  cupric  oxide.  Should  cuprous 
oxide  be  formed,  the  amount  of  extractable  copper  will  be  greatly 
reduced.  If,  for  instance,  the  extractable  copper  before  raising 
the  temperature  was  97  per  cent.,  after  an  excessive  heat  it  will 
be  reduced  to  92  or  even  90  per  cent.  When  the  roasting  is 
done  in  a  common  reverberatory  furnace,  a  mistake  of  this  kind 
can  be  corrected  by  keeping  the  charge  longer  in  the  furnace  and 
thus  oxidizing  the  cuprous  to  cupric  oxide.  In  a  mechanical  con- 
tinuously discharging  furnace,  like  the  Pearce,  however,  this  can- 
not be  done,  but  with  experience  and  care  the  decomposition  of 
the  cupric  sulphate  can  be  avoided.  It  is  of  great  importance 
that,  during  the  whole  time  of  roasting,  air  has  free  access  into 
the  furnace. 

It  is  not  possible  to  avoid  the  formation  of  some  lumps,  espe- 
cially in  roasting  leady  matte,  but  if  the  roasting  is  conducted 
properly,  these  will  be  small,  soft,  porous,  and  consist  of  well 
roasted  material.  Roasted  matte  is  always  of  coarser  grain  than 
the  raw  pulp,  and  for  this  reason  as  well  as  on  account  of  the 
lumps  it  is  necessary  to  pulverize  the  roasted  material  before 
treatment  with  sulphuric  acid.  This  is  best  done  in  a  Krupp 
ball-mill  through  a  screen  with  50  meshes  to  the  lineal  inch. 

The  dissolving  at  Argentine,  Kansas,  is  done  in  eight  agitating 
or  stir  tanks,  which  are  arranged  in  two  parallel  rows  with  a 
track  between  for  the  delivery  of  the  matte.  The  cars  are  scoop- 
shaped  and  are  partly  covered  with  sheet  iron,  and  so  made  that 
the  cover  and  scoop  end  come  close  together,  leaving  only  a  nar- 
row slit  open,  so  that  when  the  car  is  tilted  the  roasted  matte 


264 


HYDROMETALLURGY  OF  SILVER 


runs  gradually  into  the  dissolving  stir  tank.  Fig.  66  represents 
the  vertical  section  of  a  stir  tank,  12  ft.  in  diameter  and  6  ft.  deep. 
The  bottom  and  sides  are  lined  with  boards  for  protection  against 
wear  from  the  friction  of  the  pulp.  In  the  center  is  suspended  a 

Hard  Maple 
/*  Be 


Bearing  Blocks 

6ilOy.p. 


FIG.  66.  — STIR  TANK,  VERTICAL  SECTION. 

heavy  wooden  shaft,  having  fastened  to  the  lower  end,  by  heavy 
brass  plates,  a  cross-beam  cut  at  both  ends  like  a  propeller-blade. 
In  the  center,  below  the  propeller  and  fastened  to  the  bottom  of 


EXTRACTION  WITH  SULPHURIC  ACID  265 

the  tank,  is  a  cone-shaped  projection  constructed  of  strong  wooden 
staves,  the  interior  of  which  is  filled  with  sand.  This  cone  forces 
the  matte,  when  the  propeller  is  in  operation,  toward  the  pe- 
riphery, where  it  is  subjected  to  the  swift  rotating  motion  of  the 
liquor,  thus  preventing  its  accumulation  in  the  central  part, 
where  the  motion  is  much  less.  Two  outlet  pipes  provided  with 
hard-lead  valves  make  connection  with  the  pressure  tank.  These 
are  placed  one  above  the  other  to  permit  the  withdrawal  of  a' 
portion  of  the  clear  liquor  if  desired,  in  which  case  the  paddle  is 
stopped  and  the  residues  allowed  to  settle.  In  the  usual  working 
of  the  tank,  however,  the  lower  outlet  pipe  only  is  used.  The 
wooden  ring  attached  to  the  rim  of  the  tank  prevents  the  splash- 
ing of  the  pulp.  The  tank  and  trestle  support  are  placed  in  a 
flat  lead  pan  to  collect  any  leakage.  The  upper  part  of  the  shaft 
is  provided  with  a  gear-wheel  to  receive  the  power. 

The  stir  tanks  are  filled  about  two-thirds  full  with  water,  the 
agitator  set  in  motion,  and  sulphuric  acid  added  until  the  liquid 
shows  about  3  per  cent.  acid.  The  matte  is  then  charged  gradually, 
and  at  the  same  time  a  stream  of  acid  is  allowed  to  flow  in  so  as 
to  maintain  the  same  acid  strength  during  charging.  In  this 
way  the  dissolving  is  accomplished  with  an  acid  strength  of  3 
per  cent,  or  less,  and  still  yields  a  strong  solution  of  cupric  sul- 
phate. As  stated  above,  it  is  preferable  to  work  with  weak  acid, 
because  much  less  iron  and  other  impurities  will  be  dissolved  than 
with  stronger  acid.  When  the  solution  has  attained  a  strength 
of  about  20  to  22  deg.  B.,  the  flow  of  sulphuric  acid  is  stopped 
and  matte  only  charged  until  the  solution  is  neutral.  Toward 
the  end  it  is  advisable  to  charge  the  matte  at  intervals,  and  to 
make  frequent  acid  tests  to  avoid  an  excess  of  matte.  The  addi- 
tion of  sulphuric  acid  to  the  water,  and  the  matte  to  the  diluted 
acid,  produces  heat,  which  aids  the  solution  of  the  cupric  oxide. 
This  heat  is  not  sufficient,  and  the  temperature  is  further  raised 
by  a  jet  of  steam.  It  is  well  to  interrupt  the  charging  of  matte 
while  neutralizing  as  soon  as  the  solution  shows  1  per  cent,  free 
acid.  If  the  agitation  is  then  continued  this  1  per  cent,  of  acid 
will  be  diminished,  but  if  in  half  an  hour  after  the  last  acid  test 
the  percentage  of  free  acid  remains  the  same,  more  matte  is 
added  until  the  solution  is  neutral;  in  this  way  a  mistake  of  add- 
ing a  large  excess  of  matte  is  avoided. 

Below  the  stir  tanks  are  the  pressure  tanks,  into  which  the 


266 


HYDROMETALLURGY   OF  SILVER 


finished  charge  is  drawn  while  the  paddle  is  in  motion.  On 
account  of  the  residues  the  pressure  tanks  are  in  an  upright 
position,  and  are  constructed  as  follows:  The  body  consists  of 
two  cylindrical  sections  4  ft.  long  each  and  4  ft.  6  in.  in  diameter, 
the  bottom  being  of  a  spherical  shape  of  2  ft.  3  in.  radius.  The  top 
is  rounding  upward  to  a  hight  of  6  in.  above  the  rim.  The  sections 
are  tightly  flanged,  with  a  rubber  gasket  between.  Four  pipe 
connections  are  made  through  the  top;  the  discharge-pipe  enter- 
ing through  the  center  passes  nearly  to  the  bottom,  and  the  filling- 


FIGS.  67,  68,  69. -CAST-IRON  PRESSURE  TANK. 
67  (top  right),  top  view.    68,  section  on  line  AB.    69,  plan  of  supporting  frame. 

pipe,  air-inlet  and  air-outlet  pipes  are  conveniently  arranged 
around  it.  By  proper  connection  with  the  filling-pipe  and  by 
the  use  of  valves,  one  pressure  tank  is  made  to  serve  four  stir 
tanks.  The  upper  cylindrical  section  is  provided  with  a  manhole, 
and  the  tanks  are  made  of  cast  iron  and  lined  first  with  lead  and 
then  with  wood  to  protect  the  lead  from  wear  by  abrasion.  The 
air  pressure  required  is  from  40  to  50  Ib.  Fig.  67  shows  the  top 
view  and  Fig.  68  a  vertical  section  of  an  upright  pressure  tank 


EXTRACTION  WITH  SULPHURIC  ACID  267 

while  Fig.  69  illustrates  the  supporting  frame.  This  tank  as 
represented  by  the  figures  is  smaller  than  those  used  at  Argentine, 
and  has  only  one  cylindrical  section  instead  of  two.  The  general 
arrangement,  however,  is  the  same.  The  air-inlet  extends  down- 
ward along  the  side,  ending  near  the  discharge-pipe,  in  order  to 
keep  the  residues  near  the  outlet  in  a  loose  condition  and  prevent 
the  pipe  from  clogging.  For  this  reason  it  is  advisable  to  have  a 
small  stream  of  air  enter  during  filling. 

The  pulp  is  forced  by  the  pressure  tank  into  a  large  filter- 
press,  of  which  there  are  five  in  use.  The  press  is  25  ft.  long  and 
has  hardwood  frames  and  plates,  and  holds,  when  filled,  5  tons  of 
residues.  The  clarified  solution  flows  from  the  press  to  lead- 
lined  storage  tanks,  from  which  it  is  elevated,  by  means  of  a 
pressure  tank,  to  the  purifying  towers.  These  towers  were  origi- 
nally made  of  20-lb.  sheet  lead  incased  in  a  cast-iron  framework, 
which,  however,  were  replaced  by  larger  ones  made  of  4-in.  Cali- 
fornia redwood.  Fig.  70  represents  a  vertical  section  of  such  a 
tower.  They  are  made  of  staves  16  ft.  long,  9  ft.  in  diameter,  and 
are  well  hooped  with  round  iron.  The  top  and  bottom  are  flat. 
The  towers  are  firmly  fastened  to  a  strong  wooden  trestle,  which 
in  turn  is  anchored  to  a  concrete  foundation.  This  construction 
is  called  for  to  guard  against  the  oscillating  movement  of  the  tank 
from  the  action  of  the  contained  solution,  which  is  set  in  violent 
motion  by  the  ascending  air.  Inside  the  tower,  about  19  in. 
above  the  bottom,  the  4-in.  lead  air  pipe  enters,  and  is  connected 
with  a  perforated  6-in.  lead  pipe,  which  extends  diametrically  to 
the  opposite  side  and  is  closed  at  the  further  end.  Outside  the 
tower  the  pipe  extends  upward  to  above  the  top  of  the  tower, 
and  thence  downward  to  the  discharge-pipe  of  an  air-compressor 
or  to  an  air-receiver,  connected  with  the  compressor.  This 
arrangement  is  necessary  to  prevent  the  solution  from  flowing 
to  the  air-pumps  if  the  latter  be  not  in  operation.  At  the  same 
level  with  the  air-inlet,  the  4-in.  discharge-pipe  enters,  which  is 
provided  with  a  hard-lead  valve  placed  close  to  the  outside  of 
the  tower.  At  about  the  same  level  enters  a  1-in.  lead  steam-pipe 
for  heating  the  charge.  The  direct  application  of  steam  does 
not  dilute  the  solution,  because  the  evaporation,  which  is  favored 
by  the  ascending  air,  fully  equalizes  the  dilution  by  condensed 
steam.  The  top  of  the  tower  is  provided  with  an  8-in.  pipe  ex- 
tending through  the  roof  for  the  escape  of  steam  and  air,  which, 


268 


HYDROMETALLURGY   OF  SILVER 


FIG.  70.— TOWER  FOR  REFINING  CUPRIC  SULPHATE  SOLUTIONS. 


EXTRACTION  WITH  SULPHURIC  ACID  269 

however,  first  are  made  to  pass  through  a  lead-lined  box  provided 
with  shelves  arranged  in  a  zigzag  manner  to  precipitate  all  par- 
ticles of  liquor  which  may  be  carried  out  by  the  current  of  air. 
A  4-in.  pipe  also  enters  the  top,  which  serves  for  filling  the  tower 
with  crude  copper  solution.  An  opening  in  the  center  is  connected 
with  a  small  hopper  which  is  filled  with  roasted  matte.  The  upper 
third  part  of  the  tower  is  provided  with  glass  tube  gages  to  con- 
trol the  filling.  The  perforations  of  the  air-pipe  are  only  on  the 
bottom  side,  to  prevent  the  inflow  of  matte  and  precipitate.  The 
steam  pipe  enters  through  the  lid  of  the  manhole,  as  the  heat  of 
the  pipe  affects  the  wood  at  its  immediate  surrounding. 

The  tower  is  filled  with  crude  solution,  leaving,  however,  suf- 
ficient room  for  the  increase  in  volume  of  the  solution,  which 
immediately  takes  place  as  soon  as  the  compressed  air  is  supplied. 
A  tower  can  be  charged  with  about  5000  gallons  of  solution,  and 
as  three  such  charges  can  be  refined  in  twenty-four  hours  the 
working  capacity  of  one  tower  is  15,000  gallons  per  day.  At 
Argentine  eight  such  towers  are  in  use.  When  the  tower  is  filled, 
steam  is  allowed  to  enter,  and  also  some  air,  to  produce  a  more 
uniform  heating  of  the  solution.  The  air  causes  the  precipitation 
of  some  basic  iron  salts,  but  I  never  succeeded  in  precipitating 
more  than  about  half  of  the  iron  the  solution  contained,  although 
the  treatment  was  extended  many  hours.  When  the  solution 
is  hot  (75  to  80  deg.  C.)  more  air  is  admitted,  and  some  roasted 
matte  from  the  hopper  is  made  to  drop  into  the  tower.  The 
violent  boiling  motion  of  the  solution  keeps  the  matte  in  suspen- 
sion, and  after  three  to  four  hours  the  solution  will  be  entirely 
free  from  iron,  arsenic,  antimony,  etc.  To  observe  and  to  regu- 
late the  progress  of  the  operation  the  solution  is  tested  from  time 
to  time  for  iron  by  taking  samples  through  a  small  cock  inserted 
in  the  side  of  the  tower.  If  between  two  tests  the  content  of 
iron  is  not  diminished  some  more  matte  is  added.  It  is  not  neces- 
sary to  test  for  other  impurities,  because  the  iron  predominates, 
and  by  the  time  all  of  it  has  been  precipitated,  no  trace  of  any 
other  will  be  found. 

The  cupric  oxide  in  presence  of  air  combines  with  the  sul- 
phuric acid  of  the  ferrous  sulphate,  forming  cupric  sulphate, 
while  the  iron  precipitates  as  oxide  —  a  decided  advantage,  as 
the  precipitant  is  converted  into  cupric  sulphate,  and  thus 
enriches  the  solution  in  copper. 


270  HYDROMETALLURGY  OF  SILVER 

The  refined  solution  leaving  the  tower  should  not  be  stronger 
than  24  to  26  deg.  B.  when  hot,  as  otherwise  it  will  cause  trouble 
in  the  filter-press,  for  reasons  above  explained.  Should  a  con- 
centration have  taken  place  in  the  tower  beyond  this,  the  solution 
should  be  diluted.  The  tower  is  discharged  into  a  special  stir 
tank  and  from  there,  by  means  of  a  pressure  tank,  forced  into  a 
filter-press.  The  clear  and  purified  liquor  is  conveyed  to  the 
evaporating  department,  while  the  residues  are  subjected  to  an 
additional  treatment.  These  tower  residues  are  of  a  grayish 
yellow  color  and  consist  principally  of  precipitated  iron,  arsenic, 
antimony  and  some  undecomposed  matte,  with  also  some  basic 
copper  sulphate.  To  remove  the  last  named  substance  the  resi- 
dues are  treated  in  a1  stir  tank  with  2.5  to  3  per  cent,  cold  acid 
solution,  which  dissolves  the  basic  copper  salt,  leaving  the  im- 
purities unaffected,  with  the  exception  of  iron,  which  is  acted  on 
very  slightly. 

The  refined  solution  does  not  contain  even  a  trace  of  silver. 
Whatever  silver  is  converted  into  sulphate  during  the  roasting 
is  precipitated  in  the  stir  tanks  by  the  ferrous  sulphate  present 
in  the  solution.  The  residues  are  washed  well  in  the  filter-press 
to  remove  all  copper  solution.  The  wash-water  is  collected 
separately,  and  used  instead  of  fresh  water  for  the  preparation 
of  a  new  charge  in  the  stir  tanks.  The  residues  contain  all  the 
silver,  gold  and  lead  of  the  roasted  matte;  they  are  of  a  dark-red 
color  and  consist  mainly  of  iron  oxide.  They  are  sent  to  the 
lead-smelting  department  for  the  extraction  of  the  precious 
metals  and  the  lead.  On  account  of  the  large  percentage  of  iron 
oxide  and  lead  the  residues  are  an  excellent  material  for  lead 
smelting.  A  large  portion  of  the  lead  of  the  matte  is  converted 
into  sulphate  by  the  roasting,  and  therefore  does  not  act  much  as  a 
consumer  of  acid.  As  the  roasted  matte  contains  20  to  25  per 
cent,  of  its  copper  as  sulphate,  the  total  consumption  of  acid  is 
much  less  than  the  equivalent  amount  contained  in  the  blue 
vitriol  produced. 

The  evaporating  department  furnishes  90,000  gallons  of 
concentrated  solution  daily.  To  supply  this  amount  an  im- 
provement over  the  old  pan  evaporator  with  under-fire  or  steam 
coils  was  requisite,  and  as  vacuum  evaporators  could  not  be 
adopted  I  finally  devised  and  introduced  an  economical  and 
effective  evaporator,  which  is  illustrated  in  Figs.  71  and  72.  The 


EXTRACTION  WITH  SULPHURIC  ACID 


271 


272  HYDROMETALLURGY  OF  SILVER 

principle  observed  in  the  construction  of  this  evaporator  is  the 
application  of  the  hot  furnace  gases  in  a  manner  by  which  almost 
a  complete  utilization  of  the  heat  contained  in  them  takes  place. 
The  apparatus  consists  of  a  flat  tank,  wooden  with  the  exception 
of  a  2-ft.  space  at  the  front  end,  which  is  made  of  steel,  so  that  the 
wood  will  not  be  in  too  close  proximity  to  the  furnace.  The 
tank  is  65  ft.  long,  12  ft.  wide  and  2  ft.  deep,  and  is  lead  lined, 
the  two  ends  having  much  heavier  lead  lining  than  the  sides  and 
bottom.  It  is  traversed  longitudinally  by  13  6-in.  heavy  lead 
pipes.  These  pipes  rest  on  bricks  which  are  properly  placed  on 
the  bottom  of  the  tank.  The  tank  rests  on  wooden  trestle-work 
of  a  proper  hight  to  correspond  with  the  hight  of  the  furnace. 
At  the  furnace  end  in  each  lead  pipe  is  inserted  a  5-in.  pipe  4  ft. 
long,  provided  at  the  outer  end  with  a  flange.  The  iron  pipe 
serves  to  protect  the  lead  pipe  from  immediate  contact  with  the 
red-hot  gases  from  the  furnace.  They  also  make  the  connections 
between  the  lead  pipes  of  the  tank  and  the  iron  pipes  of  the 
furnace. 

The  furnace  is  comprised  of  the  fireplace  C,  the  dust-chamber 
D  and  the  distributing  chamber  D1.  At  a  proper  hight  in  the 
wall  of  the  distributing  chamber  nearest  the  tank  are  13  open- 
ings, in  each  of  which  is  inserted  a  short  cast-iron  pipe,  5  in.  in 
diameter,  with  a  flange  at  the  outer  end.  Each  pipe  is  connected 
with  its  corresponding  piece  of  iron  pipe  inserted  in  the  lead  pipe 
of  the  tank  by  a  cast-iron  S-shaped  elbow,  Y,  which  allows  the 
introduction  of  the  compressed-air  pipe  E  for  removing  any 
accumulation  of  ashes  in  the  lead  pipes  of  the  tank. 

An  American  "underfed  stoker"  is  used  for  the  slack-coal  fuel 
and  affords  practically  perfect  combustion,  which  is  of  great 
importance,  as  otherwise  the  lead  pipes  would  soon  become 
coated  with  soot  and  lose  much  of  their  efficiency  to  transmit  the 
heat  to  the  solution. 

The  opposite  ends  of  the  lead  pipes  in  the  tank  are  connected 
with  the  brick  suction  chamber  0,  which  in  turn  is  connected  by 
a  galvanized  iron  pipe,  R,  with  a  suction  fan,  P,  the  gases  being 
discharged  into  an  underground  flue,  S.  This  flue  serves  in  com- 
mon to  collect  the  waste  gases  from  11  evaporators,  and  terminates 
outside  the  building  in  a  brick  chimney  40  ft.  in  hight. 

The  top  of  the  pan  is  closed  with  a  wooden  cover,  and  wooden 
joists,  G,  are  placed  across  the  pan  about  5  ft.  apart,  having 


EXTRACTION   WITH  SULPHURIC  ACID  273 

cleats  fastened  to  the  lower  side,  as  shown  in  Fig.  71.  The  spaces 
between  the  joists  are  covered  with  boards  resting  on  the  cleats 
and  pushed  closely  together,  but  not  nailed,  so  that  the  whole  or 
part  of  the  cover  can  be  easily  removed.  About  14  ft.  from  the 
front  end  of  the  tank  is  a  14-in.  suction  pipe,  L,  connected  with 
the  main  suction  pipe,  M,  which  crosses  all  of  the  evaporators  to 
remove  the  water  vapors.  The  main  suction  pipe,  M,  as  well  as 
the  branch  pipes,  L,  are  made  of  wooden  staves  kept  tight  by 
hoops.  M  is  connected  with  a  large  suction  fan  having  the  hous- 
ing and  wings  of  sheet  copper  and  the  shaft  and  arms  of  brass. 
This  fan  rapidly  removes  the  vapor  from  each  evaporating  tank, 
and  by  its  use  the  building,  even  in  cold  winter  weather,  and  not- 
withstanding that  1 1  such  evaporators  are  in  operation,  is  entirely 
free  from  steam.  A  wooden  stack  outside  the  building  serves 
for  the  discharge  of  the  fan,  and  the  exhaust  at  each  individual 
evaporator  is  regulated  by  a  wooden  slide,  N,  inserted  below 
the  suction  pipe  L. 

Fig.  71  gives  the  construction  of  the  first  or  experimental 
evaporator.  During  the  experiments  it  was  found  that  the  6-in. 
lead  pipes  passing  through  both  ends  of  the  tank,  and  being 
burnt  with  lead  tight  to  both  the  ends,  did  not  keep  their  straight 
position,  but  on  account  of  the  expansion  became  wavy.  Profit- 
ing, by  this  experience  the  back  end  of  the  other  evaporators 
were  made  sloping,  and  the  pipes,  instead  of  passing  through, 
passed  over  the  edge  of  the  back  end.  This  allowed  the  pipes  to 
expand  freely,  and  they  retained  their  straight  position. 

Close  to  the  end  of  the  evaporating  tank  and  resting  on  the 
cover  is  the  lead-lined  feed-box,  V,  from  the  bottom  of  which  is 
a  short  pipe  or  nipple  extending  into  the  tank.  The  solution 
supply  pipe  T,  which  crosses  all  1 1  evaporators,  is  connected  with 
the  large  supply  tanks,  and  serves  to  convey  the  solution  to  each 
evaporator  by  down-takes,  U.  The  outlet  of  the  evaporating 
tank  is  in  the  side  near  the  furnace  end,  about  four  inches  above 
the  hot-air  pipes,  B  (Fig.  71). 

The  operation  is  conducted  in  the  following  manner:  The  pan 
is  first  filled  to  the  level  of  the  outlet  with  the  copper  sulphate 
solution  to  be  concentrated,  the  fire  is  then  started  and  the  stoker 
and  suction  fans  set  in  motion.  The  big  copper  fan  is  not  started 
until  the  solution  becomes  hot  enough  to  generate  steam.  Some 
solution  is  added  through  V  to  keep  the  surface  of  the  solution  at 


274  HYDROMETALLURGY  OF  SILVER 

the  same  level.  When  it  is  found  that  the  solution  near  the  out- 
let has  attained  the  desired  concentration,  a  continuous  stream  of 
refined  solution  is  allowed  to  flow  into  the  tank  from  the  feed- 
box,  which  starts  a  continuous  outflow  of  concentrated  solution 
through  the  outlet.  The  amount  of  influx  is  regulated  by  fre- 
quent hydrometer  tests  of  the  solution  at  the  outlet.  The  supply 
of  fuel  by  the  automatic  stoker  being  regular,  the  heat  of  the 
evaporator  is  very  uniform,  and  once  having  adjusted  the  proper 
influx  of  the  weak  solution  the  outflowing  stream  will  be  found 
of  quite  constant  concentration. 

The  glowing  hot  gases  entering  the  tubes  give  off  the  main 
part  of  their  heat  to  the  solution  within  a  comparatively  short 
distance  from  the  point  of  entrance,  and  cause  this  portion  of 
the  solution  to  boil.  In  the  passage  of  the  gases  through  the 
tubes  they  gradually  come  into  cooler  regions,  and  are  offered  an 
excellent  opportunity  to  give  off  more  of  their  heat  to  the  sur- 
rounding solution,  so  that,  when  they  finally  leave  the  tubes, 
their  temperature  is  much  below  the  boiling-point  of  the  solution; 
in  fact,  so  low  that  the  pipes  at  that  end  can  be  comfortably 
touched  with  the  hand.  In  a  tank  100  ft.  or  125  ft.  long  the 
gases  would  leave  at  a  temperature  about  that  of  the  surround- 
ing air,  thus  completely  utilizing  the  heat  of  the  gases. 

The  greater  economy  and  efficiency  of  this  type  of  evapora- 
tor, as  compared  with  one  having  a  steam  coil  or  bottom  fire,  is 
apparent.  The  production  of  steam  involves  a  considerable 
waste  of  heat,  and  in  using  it  to  evaporate  liquids,  its  circulation 
through  coils  produces  a  large  amount  of  condensed  water  at  a 
temperature  very  nearly  100  deg.  C.,  the  heat  of  which  is  gen- 
erally lost.  To  evaporate  by  direct  fire  under  the  bottom  of  a 
pan  is  very  inefficient  and  wasteful;  and  in  the  present  case,  in 
which  no  other  metal  but  lead  can  be  used  for  the  pan,  it  requires 
great  care  and  watchfulness  to  avoid  melting  the  metal. 

For  concentrating  chemical  solutions  which  do  not  affect  iron  the 
evaporator  described  can  be  constructed  entirely  of  iron  or  steel, 
and  at  much  less  cost.  The  pan  itself,  however,  should  always 
be  placed  in  a  wooden  tank  to  prevent  loss  of  heat  by  radiation. 

Very  fine  particles  of  ashes  settle  in  the  longitudinal  tubes, 
but  these  ashes  are  very  light,  and  by  turning  the  valve  of  the 
compressed-air  pipe  E,  one  of  which  is  attached  to  each  tube, 
they  are  easily  removed  and  blown  into  the  chamber  0. 


EXTRACTION  WITH  SULPHURIC  ACID  275 

The  continually  outflowing  concentrated  solution  passes  into 
a  lead  pipe  common  to  all  the  1 1  evaporators  and  is  conveyed  into 
a  collecting  tank,  from  which  the  liquor  is  elevated  by  means  of  a 
horizontal  pressure  tank  into  a  system  of  troughs  which  pass 
over  all  the  crystallizing  tanks,  of  which  there  are  112,  each  of 
720  cu.  ft.  capacity.  The  troughs  are  covered  and  so  arranged 
that  any  individual  tank  can  be  filled.  Wooden  tanks,  lead- 
lined,  of  the  same  size,  did  not  answer.  By  the  frequent  changes 
of  the  temperature  to  which  they  were  exposed,  the  lead  lining 
continued  to  expand  without  contracting  again,  which  caused  in 
course  of  time  so  many  leakages  that  it  became  intolerable. 
The  present  tanks  are  made  of  20-in.  thick  concrete  walls,  which, 
however,  do  not  come  up  to  expectation  either,  though  they  are 
far  superior  to  the  wooden  lead-lined  tanks.  By  the  sudden 
change  in  the  temperature  when  the  tank  is  filled  with  such  a 
large  volume  of  hot  liquor,  fine  cracks  in  the  walls  are  caused, 
through  which,  if  not  attended  to,  leakage  will  take  place,  but 
leakage  can  be  prevented  by  plastering  a  little  cement  on  the 
outside  of  the  tank.  An  experimental  tank  built  of  bricks,  how- 
ever, answered  the  requirements  of  such  large  crystallizing  tanks. 
The  brick  walls  have  in  the  center  a  2-in.  space  filled  with  a 
mixture  of  asphaltum  and  sand,  which  combines  with  the  bottom 
layer  of  the  tank,  thus  practically  forming  a  tank  by  itself,  embed- 
ded in  the  brick  work.  When  heated  by  the  sudden  filling  of  the 
tank  with  hot  solution,  the  asphaltum  softens,  and  when  gradu- 
ally cooled  contracts  without  cracking. 

On  the  top  of  each  tank  are  movable  wooden  frames  support- 
ing numerous  strips  of  lead  5  ft.  long,  on  which  the  crystals 
form,  as  well  as  on  the  sides  and  bottom.  The  solution  remains 
seven  days  in  the  tank.  In  discharging,  the  mother  liquor  is 
drawn  off  through  a  brass  tube  near  the  bottom,  then  the  frames 
with  the  strips  are  lifted  up  by  means  of  block  and  tackle  at- 
tached to  an  overhead  crawl.  The  crystals  are  knocked  off  from 
the  strips  with  a  wooden  paddle  and  fall  into  the  tank.  The 
frames  are  then  moved  to  one  side  by  means  of  the  crawl.  The 
crystals  from  the  sides  are  broken  down  also.  The  tanks  are 
arranged  in  long  rows  intersected  by  several  cross  passages. 
Between  each  two  rows  is  a  track  for  transporting  the  blue  vitriol 
crystals,  on  either  side  of  which  is  a  cement  channel  to  receive 
and  to  convey  the  mother  liquor  to  pressure  tanks  for  further 


276 


HYDROMETALLURGY  OF  SILVER 


handling.  Rails  are  laid  in  recesses  near  the  rim  of  the  tanks, 
so  that  the  rails  of  the  two  opposite  rows  of  tanks  form  a  trac"k 
for  a  hopper  mounted  on  wheels.  The  crystals  are  shoveled  with 
copper  shovels  into  this  hopper,  which  fills  the  push-car  under- 
neath through  a  spout  with  slide  in  the  center.  As  the  tanks  are 
6  ft.  deep  the  crystals  have  to  be  thrown  at  least  7  ft.,  and  in 
doing  so  some  of  the  crystals  unavoidably  fall  back  into  the 
tanks,  frequently  striking  the  shoveler.  It  was  found  that  this 
was  very  injurious  to  the  men,  especially  in  the  summer.  Their 
bodies  became  covered  with  deep  sores,  which,  while  not  danger- 
ous, were  very  painful. 


FIG.  73.  — DEVICE  FOR  DISCHARGING  BLUE  VITRIOL. 

To  protect  the  men,  I  constructed  and  set  in  operation 
the  device  illustrated  in  Fig.  73.  On  top  of  the  movable  hopper 
is  mounted  the  frame  K,  K,  with  a  turntable  and  circular  track 
underneath,  which  rests  on  a  number  of  stationary  wheels  and  is 
kept  in  place  by  the  pin  P.  The  object  of  this  turntable  is  to 
make  the  apparatus  available  for  the  opposite  row  of  tanks  by 
rotation  through  180  deg.  The  frame  K  is  provided  with  a  belt 
elevator  E,  with  copper  cups.-  On  the  shaft  F  are  the  pulleys 
L  and  M,  which  drive  the  elevator  pulley  N.  The  elevator  can 
be  brought  to  a  horizontal  position  by  means  of  the  shaft  F.  The 
lower  end  of  the  elevator  is  provided  with  the  boot  B,  which  can 
be  brought  down  to  within  a  short  distance  from  the  bottom  of 


EXTRACTION  WITH  SULPHURIC  ACID  277 

the  tank  and  into  which  the  crystals  are  shoveled.  The  power  is 
imparted  by  an  electric  motor  on  the  platform  of  the  frame, 
which  receives  the  electric  current  by  an  overhead  wire  and 
trolley.  From  the  time  this  device  came  into  operation  the 
men  were  protected,  as  they  had  to  shovel  the  crystals  only  a 
short  distance  above  the  floor. 

The  crystals  are  washed  in  a  stream  of  mother  liquor  in  a 
trough,  and  conveyed  by  this  stream  into  a  hexagonal  revolving 
screen,  having  shaft  and  arms  of  brass.  The  screen  itself  is  of 
maple  wood,  perforated.  There  are  two  such  screens,  to  make 
two  sizes  of  crystals.  The  mother  liquor  with  the  smallest  crys- 
tals, dirt  and  sediment,  after  leaving  the  second  screen,  is  con- 
veyed to  an  agitating  tank  and  heated  by  a  steam  jet  to  dissolve 
the  very  fine  crystals.  The  resultant  solution  is  sent  through 
a  filter-press,  the  clean  liquor  flowing  to  the  storage  tanks.  The 
crystals  are  dried  in  ten  brass  centrifugal  machines.  The  yearly 
production  is  about  18,000  tons  of  blue  vitriol. 

The  crystals  having  been  obtained  from  such  a  pure  neutral 
solution  are  of  a  very  deep  blue  permanent  color,  which  is  not 
affected  by  light,  except  in  the  direct  rays  of  the  sun.  They  do 
not  change  into  a  bluish  white  powder,  which  is  the  case  with 
crystals  made  from  an  acid  solution. 

The  trade  in  blue  vitriol  demands  large  crystals.  In  crystal- 
lizing a  salt  solution  in  large  tanks  it  will  be  found  that,  while  on 
the  strips  and  sides  large  crystals  are  formed,  the  bottom  will  con- 
tain mostly  small  crystals.  As  this  necessitated  the  dissolving 
and  recrystallization  of  a  large  portion  of  the  bottom  crystals, 
and  consequently  of  quite  a  percentage  of  the  total  production, 
this  peculiarity  was  rather  annoying,  and,  searching  for  the  cause, 
I  observed  that  on  the  surface  of  the  cooling  liquor  numerous 
very  small  crystals  were  formed.  They  do  not  remain  on  the 
surface,  but  sink  as  soon  as  they  are  formed.  These  crystals  are 
so  small  that  they  cannot  be  seen  as  such,  but  if  a  good  light 
strikes  the  surface,  the  liquor  right  under  the  surface  sparkles 
from  light  reflected  on  these  minute  crystals.  It  can  easily  be 
observed  that  "hey  sink  and  that  new  ones  are  continually  formed, 
thus  producing  a  very  shower  of  fine  crystals  from  the  surface 
to  the  bottom.  This  phenomenon  is  caused  by  the  evaporation  of 
the  water  on  the  very  surface  where  the  liquor  is  in  contact  with 
the  air.  By  losing  part  of  its  water  a  very  thin  sheet  of  the 


278  HYDROMETALLURGY  OF  SILVER 

solution  on  the  very  surface  will  become  so  concentrated  that  it 
has  to  form  and  drop  these  very  fine  crystals,  aided  by  the  cool- 
ing effect  of  evaporation  and  contact  with  the  air.  Following 
up  these  observations  I  caused  a  stream  of  water  to  enter  the 
tank  under  light  pressure  and  level  with  the  surface  through  a 
flat  muzzle,  so  that  the  water  did  not  mix  with  the  solution,  but 
covered  as  such  the  whole  surface  about  an  inch  thick.  This 
stopped  the  sparkling  of  these  minute  crystals  entirely,  and  when 
crystallization  was  finished  it  was  found  that  the  bottom  crystals 
were  just  as  good  as  the  crystals  on  the  sides,  so  that  they  could 
be  mixed  and  treated  together  with  the  other  crystals,  and  the 
tedious  and  expensive  operation  of  dissolving  and  recrystallizing 
of  an  already  finished  product  was  entirely  avoided.  After  this 
successful  trial  the  whole  crystallizing  plant  was  equipped  with 
a  proper  arrangement  for  this  purpose,  so  that  the  operator  had 
only  to  fill  each  tank  to  the  given  mark  and  then  to  turn  on 
the  water  for  a  short  time.  The  mother  liquor  is  somewhat  di- 
luted by  this  method,  but  the  advantage  gained  greatly  outweighs 
this  disadvantage. 

(2)    EXTRACTION  OF  SILVER  FROM  BLACK  COPPER 

This  process  is  based  on  the  reaction  which  takes  place  if 
copper  is  moistened  with  warm  diluted  sulphuric  acid  in  presence 
of  air;  the  latter  will  oxidize  the  copper,  and  the  acid  will  combine 
with  the  oxide,  forming  cupric  sulphate. 

This  method  originated  and  is  still  in  operation  at  Oker, 
Germany,  for  which  reason  it  is  also  called  the  Oker  process. 

The  material  subjected  to  this  process  is  argentiferous  and 
auriferous  black  copper.  This  is  smelted  in  a  reverberatory 
furnace,  being  partly  refined  and  then  granulated. 

The  granulated  copper  is  charged  into  wooden  lead-lined  tubs 
5  ft.  high,  with  a  bottom  diameter  of  3J  ft.  and  a  rim  diameter  of 
2J  ft.  About  5  in.  above  the  bottom  there  is  a  movable  wooden 
filter  bottom.  Below  this  bottom  an  opening  is  cut  out  4  in. 
high  and  8  in.  wide,  to  which  is  attached  a  lead  trough.  This 
opening  is  cut  so  large  because  it  serves  not  only  as  outlet 
for  the  solution,  but  acts  also  as  an  inlet  for  air.  The  boards 
forming  the  filter  bottom  are  perforated  with  inch  holes.  On 
top  of  the  filter  bottom  large  pieces  of  copper  are  placed  first, 
then  about  one  ton  of  granulated  copper  is  charged  on  top  of 


EXTRACTION  WITH  SULPHURIC  ACID  279 

the  coarse  pieces.  This  done,  a  spray  of  hot  diluted  sulphuric 
acid  of  28  deg.  B.  (heated  to  70  deg.  C.)  is  made  to  play  over 
the  copper  granules;  this  has  to  be  done  over  the  whole  surface. 
The  first  acid  charge  will  leave  the  dissolving  tank  colorless,  with- 
out containing  much,  if  any,  cupric  sulphate.  The  spray  of  acid 
is  applied  only  for  a  short  time.  By  the  evaporation  of  the  warm 
solution  and  condensation  of  the  steam  a  gentle  draft  is  produced 
through  the  opening  under  the  filter  bottom  to  the  top  of  the 
tub,  and  the  oxidation  of  the  copper  and  the  formation  of  cupric 
sulphate  takes  place.  After  a  quarter  of  an  hour  or  so  some  more 
hot  acid  is  sprayed  over  the  granules.  The  stream  of  acid  dis- 
solves the  cupric  sulphate  that  is  formed,  leaving  the  surface  of  the 
granules  clean.  The  entering  air  oxidizes  again  the  surface  of  the 
copper,  which  is  again  washed  out  by  acid.  This  is  repeated  over 
and  over  again,  and  in  fact  constitutes  the  process.  Instead  of 
pure  acid,  crude  copper  solution,  or  acid  mother  liquor,  is  mixed 
with  acid  instead  of  water,  which  has  a  much  more  energetic  dis- 
solving action  on  the  copper.  The  strength  of  such  a  solution  is 
regulated  at  32  to  34  deg.  B. 

This  is  undoubtedly  a  very  slow  and  tedious  process.  The 
oxidizing  part  of  the  process  would  be  much  hastened  if,  instead 
of  depending  on  the  natural  draft  of  a  tub  5  ft.  high,  an  artificial 
gentle  stream  of  warm  air  could  be  forced  through  the  tub,  which 
instead  of  5  ft.  in  hight  might  be  made  15  or  16  ft.  high  and 
charged  with  three  to  four  times  the  amount  of  granulated  copper. 
From  time  to  time  the  tower  could  be  flushed  with  cupric 
sulphate  solution  to  prevent  the  granules  from  being  clogged  by 
the  residue  slimes. 

Silver,  gold,  antimony,  arsenic  (the  last  coming  mostly 
from  the  crude  acid)  and  lead  remain  as  slimy  residues  which  are 
brought  out  from  the  dissolving  vat  with  every  charge  of  acid, 
and  settle  in  a  horizontal  trough  160  ft.  long,  30  in.  wide  and  14 
in.  deep.  The  solution  in  this  horizontal  trough  moves  but 
slowly,  and  when  it  reaches  the  outlet  has  cooled  to  a  tempera- 
ture only  about  2  deg.  above  the  temperature  of  the  surrounding 
air.  The  crystals  of  blue  vitriol  which  form  in  the  trough  are 
shoveled  out  on  an  inclined  bench  which  lets  the  adhering  solu- 
tion drop  back  into  the  trough.  This  is  done  every  three  days. 
The  mother  liquor  flows  into  a  pressure  tank  made  of  3-in.  wooden 
staves  and  lined  with  1-in.  lead,  which  is  well  hooped.  Instead  of 


280  HYDROMETALLURGY  OF  SILVER 

compressed  air  steam  is  used,  and  the  liquor  is  lifted  to  a  reservoir 
which  is  placed  on  a  floor  above  the  dissolving  tanks.  This 
liquor  contains  much  acid,  and  after  mixing  it  with  still  more 
acid  it  is  heated  and  used  for  dissolving  the  granulated  copper. 

To  the  crude  vitriol  which  was  shoveled  out  from  the  trough 
some  silver  slimes  are  adhering..-  They  are  first  washed  with  a 
spray  of  water,  then  dissolved  in  a  lead  pan  by  boiling  with  water 
or  with  a  mixture  of  water  and  mother  liquor  from  the  second 
crystallization.  Crude  vitriol  is  added  until  the  solution  measures 
29  deg.  B.  Then  the  fire  is  stopped  and  the  sediment  allowed  to 
settle.  When  the  temperature  has  cooled  somewhat  and  the 
solution  become  clear,  it  is  very  carefully  decanted,  so  as  not  to 
draw  off  any  of  the  silver  slimes,  and  conveyed  to  wooden  lead- 
lined  crystallizing  tanks  of  147  cu.  ft.  capacity.  Ten  days  is 
allowed  for  crystallization. 

The  silver  slimes  consist  of: 

Silver 3.068  per  cent. 

Metallic  copper 7.400  per  cent. 

Lead 23.100  per  cent. 

Lime 8.300  per  cent. 

Sulphuric  acid    16.200  per  cent. 

Antimony  and  arsenic     27.000  per  cent. 

The  large  percentage  of  antimony  and  arsenic  comes  from 
the  impure  chamber  acid  which  is  used  for  dissolving  the  copper. 
These  slimes  are  well  washed  and  mixed  with  litharge  while  still 
wet,  then  dried  and  melted;  the  resulting  lead  contains  1£  per 
cent,  silver. 


XXI 

THE  ZIERVOGEL  PROCESS 

IN  this  process  hot  water  is  used  as  solvent  for  the  silver, 
and  the  latter  has  to  be  first  converted  into  sulphate,  while  the 
other  metal  sulphides,  copper  and  iron,  have  to  be  converted  into 
oxides.  The  only  suitable  material  is  copper  matte  containing  a 
certain  percentage  of  iron.  If  the  matte  contains  too  much  or 
too  little  iron  the  extraction  of  silver  becomes  inferior.  Really 
good  results  were  only  obtained  from  the  copper  matte  of  Mans- 
feld,  which  contains  80  percent,  copper  sulphide,  11  per  cent,  iron 
sulphide  and  0.4  per  cent,  silver. 

The  great  sensitiveness  of  this  process,  and  the  complicated 
roasting  it  requires,  limit  its  application.  The  roasting  was 
described  in  Chapter  X,  on  sulphating  roasting,  and  we  have, 
therefore,  only  to  consider  the  extraction  of  the  silver  from  the 
roasted  matte.  The  roasted  material  is  charged  in  lots  of  500 
Ib.  in  small  tubs  25  in.  in  diameter  and  24  in.  high.  ^These  tubs 
are  provided  with  a  filter  bottom,  and  a  number  of  them  are  placed 
in  one  row.  The  leaching  is  done  with  hot  water  of  about  85  deg. 
C.  The  water  passing  through  the  roasted  matte  dissolves  the 
silver  sulphate,  and  leaves  the  tub  through  an  outlet  pipe  below 
the  filter  bottom.  It  flows  into  a  square  settling-tank,  which 
extends  the  whole  length  of  the  row  and  which  receives  the  silver 
solution  from  all  the  leaching-tubs.  This  tank  is  30  ft.  long,  2  ft. 
wide  and  1J  ft.  high,  and  has  longitudinally  a  partition  wiiich  is 
lower  than  the  rim  of  the  tank,  and  over  which  the  silver  solu- 
tion flows  into  the  second  half  of  the  tank  when  the  first  one  is 
filled.  This  tank  serves  to  catch  any  material  that  may  be  car- 
ried out  by  the  stream  from  the  leaching-tubs.  Below  this  set- 
tling-tank is  placed  a  row  of  tubs  21  in.  in  diameter  and  20  in. 
deep,  corresponding  in  number  to  that  of  the  leaching-tubs.  Each 
of  them  is  provided  with  a  filter  bottom.  They  serve  for  precipi- 

281 


282  HYDROMETALLURGY   OF  SILVER 

tating  the  silver  from  the  solution,  and  are  filled  with  copper 
bars  1  in.  thick,  5  in.  wide  and  14  in.  long.  In  these  tubs  nearly 
all  the  silver  is  precipitated  as  cement  silver.  Flowing  out  from 
under  the  filter  bottom  the  solution  enters  a  lead-lined  box  15  in. 
wide  and  6  in.  deep,  which  extends  in  front  of  all  the  precipitation 
filters.  The  bottom  of  this  box,  which  has  several  outlets,  is 
covered  with  small  pieces  of  copper.  Under  each  outlet  is  placed 
a  tub  with  filter  bottom  on  which  bar  copper  and  granulated 
copper  are  placed,  and  through  which  the  solution  from  the  out- 
lets passes.  In  these  tubs  but  very  little  silver  is  precipitated,  and 
the  solution  when  leaving  them  is  free  from  silver.  If  the  water 
used  for  dissolving  the  silver  from  the  roasted  matte  is  slightly 
acidified  it  hastens  the  dissolving  and  causes  better  extraction; 
it  prevents  the  separation  of  basic  salts. 

The  cement  silver  contains  some  metallic  copper  and  some 
gypsum.  For  purification,  it  is  placed  in  tubs  and  rubbed  with 
wooden  pestles  to  free  the  copper  from  the  adhering  silver,  then 
washed  in  hand  pans  to  separate  the  coarser  copper,  next  placed 
in  tubs  and  digested  with  diluted  sulphuric  acid  during  six  to 
seven  days  to  remove  the  copper  and  as  much  as  possible  of  the 
gypsum,  and  finally  it  is  washed  with  hot  water.  The  washed 
cement  silver  contains  in  1000  parts  860  to  870  parts  of  fine  silver. 
It  is  dried  and  smelted.  The  desilverized  solution  is  made  to  pass 
over  scrap  iron  to  precipitate  any  copper  that  may  have  dis- 
solved from  the  matte  by  the  acidified  water.  Usually,  however, 
it  is  mixed  with  water  and  used  over  again  a  number  of  times 
to  dissolve  silver  instead  of  using  pure  water  for  the  latter 
purpose. 


XXII 

TREATMENT   OF   SILVER  ORES  RICH  IN   GOLD 

To  me  was  given  the  task  of  working  the  rich  silver-  and  gold- 
bearing  concentrates  from  the  old  Tarshish  mine,  Alpine  county, 
California.  These  concentrated  sulphides  contained  258  oz. 
silver  and  over  10  oz.  in  gold  per  ton,  which  represents  the  average 
assay  of  five  months.  All  the  gold  was  contained  in  the  sulphurets; 
no  metallic  gold  could  be  detected.  The  plan  of  operation 
was  self-suggesting,  viz.:  to  subject  the  material  to  a  chlori- 
dizing  roasting,  then  to  impregnate  it  by  Plattner's  method  with 
chlorine,  and  then  leach,  first  with  water  for  the  extraction  of  the 
gold,  and  afterward  with  sodium  hyposulphite  for  the  extraction 
of  the  silver.  When  this  scheme  was  executed  it  was  found,  how- 
ever, that  while  a  high  percentage  of  silver  could  be  extracted,  only 
about  50  per  cent,  of  the  gold  was  thus  extractable.  The  reason 
is  not  easily  explained.  After  operating  for  a  while  in  this  way 
and  paying  particular  attention  to  the  roasting  without  obtaining 
any  better  result,  the  operations  of  the  process  were  reversed. 
The  ore,  after  being  roasted  with  salt,  was  leached  first  with  water 
and  with  sodium  hyposulphite  for  the  extraction  of  the  silver, 
and  then  treated  with  chlorine  gas  by  Plattner's  method.  This 
reversing  of  the  operations  had  a  most  beneficial  influence  on  the 
result,  effecting  an  extraction  of  95  per  cent,  of  the  gold. 

The  operations  were  as  follows: 

(1)  Roasting.  —  The  roasting  was  done  with  10  per  cent,  of 
salt,  the  salt  being  added  after  the  oxidation  had  progressed  for 
some  time.  Soon  after  the  salt  was  added  free  gold  could  be 
detected  by  concentrating  a  sample  of  the  ore  in  a  horn  spoon. 
Roasting  was  continued  until  the  concentration  test  did  not  show 
any  undecomposed  sulphurets  and  but  very  little  magnetic  iron, 
when  the  concentrated  part  of  the  sample  was  tested  with  a  mag- 
net, while  a  large  amount  of  very  bright  yellow  gold  was  visible. 

283 


284  HYDROMETALLURGY  OF  SILVER 

The  roasted  charge,  after  cooling,  was  sifted  through  a  screen 
of  10  meshes  to  the  running  inch.  The  fine  was  charged  into  the 
filter  tanks  while  the  coarse  was  accumulated  in  a  larger  lot,  then 
crushed  dry  in  a  battery  and  slightly  roasted. 

(2)  Base-Metal  Leaching.  —  In  order  to  prevent,  or  rather  to 
greatly  diminish,  the  dissolving  of  silver  chloride  by  the  base- 
metal  solution,  I  devised  and  practised  the  following  mode  of 
operation:  Instead  of  applying  the  stream  of  water  on  top  of  the 
ore,  as  is  usually  done,  the  water  was  made,  under  very  slight 
pressure,  to  enter  the  vat  under  the  filter  bottom  and  to  ascend 
gradually  through  the  ore.     In  this  way  the  concentrated  part  of 
the  solution  which  dissolves  the  silver  chloride  accumulated  on 
top  of  the  ore.     If,  then,  this  concentrated  solution  was  diluted  by 
a  stream  of  water  applied  on  top  and  the  solution  was  permitted  to 
flow  out  from  under  the  filter  bottom,  the  silver  chloride  was 
precipitated  on  and  through  the  ore,  and  was  dissolved  again  by 
the  subsequent  leaching  with  sodium  hyposulphite.     Sufficient 
room  in  the  vat  was  left  above  the  ore  for  this  operation. 

(3)  Leaching  the  Silver.  —  This  was  done  in  the  usual  way  by 
applying  a  diluted  solution  of  sodium  hyposulphite.     The  result- 
ing silver  bullion  was  957  fine. 

(4)  Second  Leaching  with  Water.  —  After  the  silver  was  ex- 
tracted the  solution  of  sodium  hyposulphite  was  pressed  out  by 
water,  and   washing  was   continued   until  the  outflowing  liquid 
was  perfectly  free  from  sodium  hyposulphite.     Such  a  careful 
washing  is  necessary,  because  sodium  hyposulphite  added  to  a 
solution  of  gold  chloride*  prevents  the  precipitation  of  the  gold 
by  ferrous  sulphate. 

The  desilverized  and  washed  ore  was  removed  from  the  vat 
to  a  drying  kiln,  where  it  was  left  for  a  time  till  the  surplus 
water  had  evaporated.  After  this  it  was  charged  back  into  the 
vat,  still  moist.  This  second  handling  and  drying  cannot  be 
avoided,  because  the  ore  after  leaching  is  too  wet  and  tightly 
packed  to  permit  a  free  passage  of  the  chlorine  gas. 

(5)  Extraction  of  the  Gold.  —  After  the  extraction  of  the  base- 
metal  chlorides  and  the  silver  the  gold  is  left  in  a  metallic  state, 
and  bright  and  clean,  permitting  a  very  close  extraction.     On  the 
inside  periphery  of  the  vat  a  groove  was  cut  into  the  staves  from 
the  rim  down,  forming  a  shoulder  or  recess  into  which  a  tight 
wooden  cover  fitted.     The  shoulder  was  2J  in.  below  the  rim,  so 


TREATMENT  OF  SILVER  ORES  RICH  IN  GOLD  285 

that  when  the  1-in.  cover  was  put  on  the  staves  projected  1 J  in. 
above  the  cover.  Around  the  periphery  the  cover  was  tightly 
luted  with  clay,  and  then  water  was  poured  on  it  to  about  the  depth 
of  one  inch.  This  sheet  of  water  kept  the  cover  perfectly  tight. 
The  water,  however,  was  not  poured  on  the  cover  until  the  gas 
appeared  on  the  surface  of  the  ore.  The  cover  was  provided  with 
two  IJ-in.  pieces  of  pipe  projecting  about  6  in.  above  the  cover, 
and  a  square  opening  6  x  6  in.  When  the  chlorine  gas  appeared 
above  the  ore  this  opening  was  closed  with  a  cover  luted  tight 
with  clay  and  the  water  poured  on  top  of  the  cover.  When  the 
gas  commenced  to  escape  through  the  pipes  in  the  cover  they 
both  were  closed  with  balls  of  clay.  The  ore  was  left  in  contact 
with  the  chlorine  gas  for  twelve  hours,  and  as  soon  as  it  was 
ready  for  the  extraction  these  clay  balls  were  removed  and  one 
of  the  pipes  was  connected  with  the  water-pipe  by  a  hose,  while  the 
other  was  connected,  by  means  of  a  hose,  either  with  another  vat 
already  prepared  for  chlorination,  or  with  the  ash-pit  of  the 
roasting  furnace.  This  was  done  to  utilize  the  surplus  of 
chlorine  gas,  and  to  protect  the  workmen  from  its  very  injurious 
effect.  Care  was  taken  to  place  a  sack,  kept  in  place  by  bricks, 
on  top  of  the  ore  right  under  the  water  inlet,  in  order  to  prevent 
the  stream  from  working  into  the  ore. 

Chlorine  was  generated  in  a  leaden  gas  generator  heated  by 
steam. 

The  gold  solution  was  collected  in  precipitation  tanks  and 
precipitated  with  a  solution  of  ferrous  sulphate.  Separate  tanks 
were  used  for  the  precipitation  of  the  silver.  There  was  also  a 
separate  line  of  troughs  for  each  metal,  to  guard  against  the  enter- 
ing of  any  sodium  sulphite  solution  into  the  gold  solution,  because, 
as  stated  above,  sodium  hyposulphite  prevents  the  precipitation 
of  the  gold  by  ferrous  sulphate. 

In  working  ordinary  gold-bearing  sulphurets  by  the  Plattner 
method,  the  gold  solution  turns  jet  black  when  the  iron  solution 
is  added,  which  is  caused  by  the  precipitation  of  the  gold  in  metal- 
lic state,  but  in  such  an  extremely  finely  divided  condition  that 
it  assumes  this  color.  The  concentrated  sulphurets  from  the 
Tarshish  mine  were  very  variable  in  their  silver  and  gold  contents, 
and  sometimes  lots  were  treated  containing  as  much  as  $700  to 
$800  per  ton  in  gold.  When  such  rich  gold  ore  is  chloridized, 
the  solution  carrying  out  the  gold  is  of  a  very  lustrous  yellow 


286  HYDROMETALLURGY  OF  SILVER 

color,  and  if  ferrous  sulphate  solution  is  added  red-brown  clouds 
are  formed,  which  rapidly  sink  to  the  bottom.  There  the  gold 
accumulates  in  spongy  lumps  of  great  specific  gravity,  and  some 
of  them  show  scales  of  bright  gold,  which  under  the  microscope 
might  prove  to  be  crystallized  gold.  There  is  but  very  little 
more  time  used  in  leaching  rich  gold  ore  than  poor,  on  account  of 
the  great  solubility  of  gold  chloride  in  water. 

The  gold  was  well  washed,  dried  and  melted  with  borax,  while 
the  silver  precipitate  was  melted  with  iron  and  borax  in  graphite 
crucibles. 

Results.  —  To  ascertain  the  working  results  of  this  method 
the  concentrated  raw  sulphurets  delivered  to  the  reduction  works 
each  day  were  carefully  weighed  and  assayed  during  a  period  of 
five  months.  The  average  value  of  these  concentrates,  as  men- 
tioned above,  was  258  oz.  silver  and  a  little  over  10  oz.  gold  per 
ton.  The  total  value  of  the  bullion  shipped  during  this  period, 
compared  with  the  value  of  the  raw  sulphurets  worked  during 
this  time,  showed  an  actual  extraction  of  silver  96  per  cent,  and 
of  gold  95  per  cent. 

The  gold  obtained  was  of  high  fineness,  varying  from 
970/1000  to  987/1000. 


XXIII 

CYANIDATION   OF   AURIFEROUS  SILVER   ORES 

THE  cyanide  process  is  based  on  the  fact  that  gold  and  silver 
in  presence  of  oxygen  dissolve  in  an  alkaline  solution  of  potas- 
sium or  sodium  cyanide.  The  cyanide  solution,  however,  dis- 
solves also  other  metals  and  their  sulphides,  like  iron,  copper, 
zinc,  lead,  antimony,  arsenic,  etc.,  but  it  dissolves  them  much 
slower,  which  condition  makes  the  process  possible,  because,  in 
dissolving,  these  substances  decompose  the  potassium  cyanide, 
and  as  they  as  a  rule  offer  a  very  much  larger  surface  than  the 
gold  and  silver,  they  would  at  equal  dissolving  ratio  consume 
so  much  potassium  cyanide  as  to  render  the  process  financially 
impracticable. 

Some  of  the  above-named  metals  and  their  compounds  act 
more  energetically  on  the  potassium  cyanide  than  do  others, 
but  as  they  all  act  deterioratingly  it  is  apparent  that  ores 
heavily  charged  with  sulphureted  minerals  are  not  suitable  for 
this  process  unless  they  are  first  subjected  to  a  chloridizing  roast- 
ing. This,  of  course,  reduces  greatly  the  applicability  of  the 
process  for  silver  ores,  because  much  the  larger  part  of  them 
are  complex  sulphureted  ores.  If  the  ore  has  to  be  roasted  with 
salt,  the  process  enters  into  competition  with  the  lixiviation 
process  with  sodium  hyposulphite,  and  then  it  is  doubtful  whether 
it  will  prove  to  be  superior,  except  when  the  ore  contains  a  suf- 
ficient amount  of  gold. 

TREATMENT  OF  RAW  ORE 

To  be  suitable  for  this  treatment  the  ore  has  to  be  but  slightly 
mineralized  and  the  silver  ore  to  occur  in  its  purer  varieties,  as 
sulphide,  chloride  or  chlorobromide.  But  even  from  such  clean 
ores  the  extraction  percentage  of  the  silver  varies  greatly.  Some 
of  them  yield  90  per  cent,  and  more  of  their  silver  to  the  solvent, 

287 


288  HYDROMETALLURGY  OF  SILVER 

while  others  of  similar  character  will  yield  only  50  per  cent. 
Much  more  reliable  and  uniform  results  are  obtained  with  regard 
to  the  extraction  of  the  gold,  for  which  reason  the  process  is  more 
suitable  for  treating  clean  auriferous  silver  ores  than  silver  ores 
which  do  not  contain  any  gold.  Large  quantities  of  ores  can  be 
cheaply  treated  by  this  process  in  simple  appliances  requiring 
but  little  repair,  which  give  the  means  of  working  low-grade  ores 
which  by  no  other  known  process  could  be  worked  profitably. 
A  number  of  mining  properties  are  worked  now  with  profit  which 
formerly  were  not  productive  because  the  ore  was  too  poor  in  silver 
and  gold  to  be  treated  by  other  more  costly  processes. 

Raw  ores  when  finely  pulverized  permit  only  a  very  slow  per- 
colation of  solution,  owing  to  exceedingly  fine  slimes  which  are 
formed  in  pulverizing,  and  which,  in  some  cases,  pack  so  as  to  be 
practically  impenetrable.  To  overcome  this  difficulty,  the  ore  is 
crushed  wet,  and  the  slimes  are  separated  from  the  sand  either  by 
cones  and  other  sand-separating  appliances  or  by  conveying  the 
pulp  from  the  stamp  battery  direct,  or,  where  concentration  is 
practised,  from  the  discharge  of  the  concentrating  tables  to  large 
masonry  tanks  to  retain  the  sand,  and  the  overflow  of  these  tanks 
to  another  series  of  tanks  to  collect  the  slimes.  The  latter  is  a 
rather  crude  method,  permitting  only  a  very  imperfect  separation, 
and  necessitates  considerable  handling  of  the  material,  while  by 
the  former  method  the  filling  of  the  tanks  is  done  automatically. 
However,  there  may  be  circumstances  which  make  the  adoption 
of  the  second  method  necessary  or  even  more  practicable. 

In  the  following  I  give  an  abstract  of  a  very  careful,  intelli- 
gent and  exhaustive  record  of  the  cyaniding  of  auriferous  silver 
ores  of  Palmarejo,  Chihuahua,  Mexico,  written  by  T.  H.  Oxnam, 
mining  engineer,  Palmarejo  &  Mexican  Gold  Fields,  Ltd./ 
Chinipas,  and  read  at  the  Washington  meeting,  May,  1905,  of 
the  American  Institute  of  Mining  Engineers: 

CYANIDING  AURIFEROUS  SILVER  ORES  AT  PALMAREJO, 
MEXICO 

The  predominating  value  of  the  ore  now  treated  by  the  Pal- 
marejo and  Mexican  Gold  Fields,  Ltd.,  is  silver.  The  method 
consists  of  wet-crushing  and  concentrating,  followed  by  cyanida- 
tion  of  the  unroasted  sand  and  slime. 

The  Palmarejo  mines  are  located  in  the  southwestern  part  of 


CYANIDATION   OF   AURIFEROUS   SILVER  ORES  289 

Chihuahua,  on  the  foothills  of  the  Sierra  Madre,  and  at  an  eleva- 
tion of  3200  ft.  The  mills,  12  miles  distant,  are  situated  on  the 
Chinipas  river,  near  the  town  of  Chinipas,  which  is  150  miles 
northeast  of  Agiabampo,  on  the  Gulf  of  California.  Supplies  are 
shipped  via  this  port. 

M ill  and  Cyanide  Plant.  —  The  50-stamp  mill  and  cyanide 
plant  are  situated  on  the  Chinipas  river,  1.5  miles  east  of  Chinipas, 
at  a  place  known  locally  as  "  El  Zapote."  Water-power,  furnished 
by  the  river,  is  used  to  run  the  mill,  slime  plant  and  machine-shop. 
A  masonry  conduit  11  miles  long  conducts  the  water  to  a  pen- 
stock a  short  distance  above  the  mill,  thence  through  a  steel 
pipe,  1100  ft.  long,  tapering  from  48  in.  in  diameter  at  the  pen- 
stock to  22  in.  at  the  wheel-pits;  here  there  are  four  6-ft.  Pelton 
wheels  under  a  97.5-ft.  head. 

The  ore  consists  essentially  of  a  silicious  matrix  in  which  is 
disseminated  a  small  percentage  of  pyrite.  Black  manganese 
oxide,  and  calcite  are  present  in  varying  proportions,  and  small 
quantities  of  antimony  and  arsenic,  together  with  traces  of  bis- 
muth, also  occur.  Occasional  traces  of  copper  and  zinc  are  found. 
The  major  portion  of  the  silver  occurs  in  the  form  of  argentite, 
though  a  certain  amount  of  stephanite  is  present,  and  occasionally 
small  patches  of  chlorobromide  and  native  silver. 

System  of  Milling.  —  The  ore,  averaging  6  per  cent,  moisture, 
is  brought  to  the  mill  in  trains  of  from  9  to  14  cars,  and  is  dumped 
into  the  main  upper  storage-bin,  which  has  a  capacity  of  1100 
tons.  From  this  the  ore  is  drawn  over  3.5  x  10-ft.  iron  grizzlies 
having  1.5-in.  openings  to  the  7  x  10-in.  Blake  rock-crushers, 
which  run  at  250  r.  p.  m.  and  crush  to  2-in.  size.  Of  the  dump- 
ore,  which  is  coarse  and  extremely  hard,  approximately  90  per 
cent,  goes  to  the  crushers;  of  the  mine-ore,  which  is  finer  and 
softer,  approximately  50  per  cent,  goes  to  the  crushers,  the  other 
10  and  50  per  cent.,  respectively,  falling  through  the  grizzlies. 

A  secondary  storage-bin  (of  1100-ton)  receives  the  ore  from 
both  grizzlies  and  crushers.  The  ore  is  then  trammed  to  three 
small  intermediate  bins,  each  of  50  tons  capacity;  from  here 
it  is  conveyed,  by  means  of  half-ton  cars,  to  the  hoppers  of  the 
Challenge  ore-feeders.  This  double  handling  of  the  ore  is  incon- 
venient, but  is  rendered  necessary  because  of  the  construction 
of  the  mill,  which  was  originally  erected  for  different  require- 
ments. 


290  HYDROMETALLURGY  OF  SILVER 

The  stamps,  when  equipped  with  new  shoes,  weigh  850  Ib. 
They  drop  6  to  7  in.,  100  times  per  min.,  the  order  of  drop  being 
1 — 3 — 5 — 2 — 4;  20-mesh  brass-wire  screen,  No.  26  wire,  is  used; 
the  hight  of  discharge  is  kept  at  2  in.  The  stamp-duty  is  from 
2.75  to  3.25  tons  per  twenty-four  hours.  The  average  stamp- 
duty  would  doubtless  be  somewhat  increased  by  the  installation 
of  narrow  mortars  of  the  Homestake  pattern.  For  some  time 
past,  forged-steel  shoes  have  been  used  in  preference  to  the  cast- 
iron  shoes  of  our  own  make.  The  steel  shoes  cost  15c.  per  ton  of 
ore  crushed,  as  compared  with  18c.  for  the  cast-iron  shoes.  We 
cast  all  our  own  dies,  for  which  purpose  the  worn-out  shoes  and 
iron  and  steel  scrap  are  employed.  The  average  life  of  a  forged- 
steel  shoe  is  95  days,  while  that  of  a  cast-iron  die  is  33  days. 

From  the  batteries  the  pulp  passes  directly  over  10  Wilfley 
concentrators,  running  with  a  |-in.  stroke  at  215  strokes  per  min. 
During  the  year  ending  July  1,  1904,  the  concentrator  saved 
0.76  per  cent,  by  weight  of  the  ore  as  a  product  which  contained 
18.28  per  cent,  of  the  gold  and  17.98  per  cent,  of  the  silver  of  the 
ore. 

A  wooden  launder  conveys  the  pulp  from  the  tables  to  the 
tailing  elevator-wheel.  The  latter  is  14  ft.  in  diameter  and  is  of 
the  outside-bucket  type,  having  22  steel  buckets  (each  18  in. 
long,  8.5  in.  wide  and  8.5  in.  deep,  with  a  capacity  of  1025  cu.  in.). 
The  wheel  is  driven  by  a  f-in.  plow-steel  wire  cable  at  a  speed  of 
18  r.  p.  m.  The  discharge  efficiency,  as  in  all  wheels  of  this  type, 
is  not  high,  the  tailing  leaving  the  wheel  in  a  launder  5.5  ft.  above 
the  level  of  the  mill  launder  supplying  the  pulp. 

A  large  masonry  sand -retaining  tank  (divided  into  four 
compartments,  each  compartment  measuring  25  x  80  x  4  ft. 
in  depth)  receives  the  product  from  the  wheel.  Distribution  is 
effected  by  a  central  launder  in  each  compartment,  provided 
with  a  number  of  4-in.  side-discharge  pipes.  Each  compartment 
is  provided  with  a  removable  end-discharge  gate,  4  ft.  wide,  com- 
posed of  pieces  of  2-in.  plank,  planed  smooth  on  the  edges  and 
sliding  in  guides  secured  to  the  side-posts.  As  the  compartment 
fills  up  with  sand,  the  discharge  of  these  gates  is  raised.  The 
discharge  overflow  empties  into  the  main  slime  launder.  Each 
compartment  also  communicates  with  its  immediate  neighbors 
by  small  side-discharge  doors.  The  purpose  of  this  arrangement 
is  that  the  mill  product  may  be  emptying  into  one  compartment, 


CYANIDATION   OF  AURIFEROUS  SILVER  ORES  291 

from  which  a  portion  of  the  finer  material  escapes  through  one 
of  the  side-gates  to  an  adjoining  compartment,  while  the  finest 
material  is  passing  off  in  the  discharge  over  the  lowered  end  gate 
of  this  second  compartment. 

It  is  found,  however,  that  a  considerable  quantity  of  the  finest 
material  will  always  tend  to  collect  at  the  lower  end  of  the  first 
compartment,  receiving  the  discharge  of  the  elevator-wheel,  to 
lessen  which  an  overflow  from  the  end  gate  of  this  compartment 
is  also  necessary.  The  first  five  or  six  tons  of  material  removed 
from  the  compartments  is  always  slimy,  and  is  trammed  a  short 
distance  to  an  open  drying-patio,  where  it  is  spread  out,  sun- 
dried  and  broken  up;  after  this  it  is  mixed  in  with  the  coarser 
sand  and  treated  in  the  leaching- vats.  A  third  compartment  of 
the  sand-retaining  tank  is  kept  full  of  sand,  which  is  being  drained 
while  the  dry  sand  is  trammed  from  the  fourth  compartment. 
Each  of  these  compartments  holds  the  sand  of  forty-eight  to  sixty 
hours'  crushing  in  the  mill.  The  retained  sand  is  usually  sub- 
jected to  two  days'  draining  before  charging  it  into  the  leaching- 
vats.  The  fine  material  escaping  in  the  overflow  from  the  masonry 
retaining-tank  is  carried  by  a  wooden  launder  to  three  so-called 
"slime  pits/'  having  an  aggregate  capacity  of  15,000  tons.  Every 
precaution  is  exercised  that  no  slime  escapes  at  the  overflow 
gates  of  these  pits;  but  at  no  time  is  such  overflow  perfectly  free 
from  suspended  matter.  During  the  18  months  (ending  Decem- 
ber 31,  1904)  of  the  total  net  tonnage  crushed  in  the  mill,  19.16 
per  cent,  went  to  the  slime  pits.  Sizing  tests  (using  the  ordinary 
brass-wire  assay  screens)  have  shown  that  about  6  per  cent,  of 
this  material  is  retained  on  the  100- mesh,  while  85  per  cent,  passes 
a  200-mesh  screen. 

Although  this  material  is  chiefly  slime  (which  on  long  drying 
cracks  up  into  layers  almost  impervious  to  leaching),  yet  it  is 
found  that  considerable  fine  but  leachable  sand  is  deposited  at 
the  head  of  the  slime  pits  and  near  the  discharge  from  the  slime 
launder.  About  two  months  after  ceasing  to  discharge  into  any 
one  slime  pit,  this  fine  sand  at  the  head  of  the  pits  dries  sufficiently 
(during  ordinary  weather)  to  permit  of  being  walked  on;  it  is 
then  conveyed,  by  contract  labor,  to  the  open  drying-floor,  or 
patio,  together  with  a  certain  percentage  of  more  slimy  material 
which  unavoidably  becomes  mixed  with  it.  Here  it  is  spread 
out,  sun-dried  and  thoroughly  broken  up,  after  which  it  is 


292  HYDROMETALLURGY  OF  SILVER 

mixed  in  with  the  ordinary  sand  and  treated  by  leaching.  Dur- 
ing the  past  year  (1904)  2400  tons  of  very  fine  material  from  the 
slime  pits  has  been  so  treated.  By  far  the  greater  portion  of  the 
material  collected  in  the  slime  pits,  however,  is  so  extremely  fine 
and  has  such  a  clayey  nature  that  it  is  almost  impervious  to  leach- 
ing. This  portion,  slime,  is  allowed  to  dry  as  much  as  practicable, 
and  is  then  treated  by  agitation  in  a  separate  plant,  as  will  be 
described  further  on  in  this  paper.  Figs.  74  and  75  show  the 
arrangement  of  the  cyanide  leaching  plant. 

Cyanidation  of  Sand.  —  The  sand  caught  in  the  large  masonry 
sand-retaining  tank  (after  being  allowed  to  drain  as  long  as  pos- 
sible, usually  from  thirty-six  to  forty-eight  hours)  is  trammed 
in  half-ton  cars  to  the  cyanide  leaching- vats;  these  are  12  in  num- 
ber, 30  ft.  in  diameter  and  4.5  ft.  deep.  The  filter  bottom  (which 
reduces  the  available  depth  to  4  ft.  2  in.)  consists  of  a  wooden, 
lattice  framework,  covered  by  a  layer  of  cocoa  matting,  over 
which  is  stretched  a  filter  cloth  of  8-oz.  duck.  Two  heavier 
grades  of  duck  have  been  tried,  but  they  reduced  the  rate  of 
leaching  and  gave  less  satisfactory  service  than  the  8-oz.  cloth; 
10  of  the  leaching- vats  are  constructed  of  No.  9  sheet  steel;  the 
other  two  were  built  on  the  premises  of  3-in.  native  pine.  Two 
additional  vats,  of  the  same  dimensions  and  capacity,  made  of 
3-in.  redwood  throughout,  are  in  course  of  erection. 

The  sand,  as  charged  into  the  leaching-vats,  carries  14  to  16 
per  cent,  of  moisture;  each  vat  holds  100  tons  of  dry  sand.  While 
being  trammed  to  the  leaching-vats,  slaked  lime  is  added  to  each 
car,  and  in  the  proportion  of  4  to  5  Ib.  of  lime  per  ton.  The  vats 
are  filled  and  discharged  by  contract,  for  $19  a  vat,  equivalent  to 
$0.19  per  ton. 

Two  stock  solutions  are  employed:  the  weak,  of  0.25  to  0.30 
per  cent.  KCN;  the  strong,  of  0.75  to  0.80  per  cent.  KCN.  The 
working  strength  of  the  solutions  is  always  taken  as  that  indicated 
by  titration  with  silver  nitrate  (in  presence  of  a  few  drops  of  a 
10  per  cent,  solution  of  potassium  iodide,  as  an  indicator),  10  c.c. 
of  the  cyanide  solution  being  taken  for  titration.  For  convenience, 
we  still  express  the  strength  of  our  working  solutions  in  terms  of 
potassium  cyanide,  although  for  over  a  year  past  we  have  been 
employing  sodium  cyanide  exclusively.  Titration  with  silver 
nitrate  shows  that  the  sodium  cyanide  used  is  equivalent  to  about 
125  per  cent,  of  potassium  cyanide.  Our  experience  with  sodium 


CYANIDATION   OF  AURIFEROUS  SILVER  ORES  293 


FIG.  74.  — CYANIDE  LEACHING  PLANT, 
PLAN. 


294 


HYDROMETALLURGY  OF  SILVER 


CYANIDATION   OF  AURIFEROUS  SILVER  ORES  295 

cyanide  leads  us  to  believe  that  it  is  fully  as  efficient  as  potassium 
cyanide.  It  also  appears  that,  since  commencing  the  exclusive 
use  of  sodium  cyanide,  our  solutions  become  less  fouled  than  pre- 
viously. By  the  adoption  of  sodium  cyanide  a  saving  of  20  per 
cent,  of  the  freighting  expense  on  this  article  has  been  effected. 
Besides  the  saving  in  transportation  expenses,  the  sodium  cyanide 
appears  to  possess  other  advantages.  Other  things  being  equal, 
it  would  seem  preferable  to  use  a  salt  as  pure  as  can  be  obtained. 
Absolutely  pure  sodium  cyanide  is  equivalent  to  about  132  per 
cent,  of  potassium  cyanide;  a  product,  testing  from  125  to  130 
per  cent,  of  potassium  cyanide,  is  nearly  pure.  It  by  no  means 
follows,  however,  that  the  ordinary  commercial  cyanide,  rated  as 
98  to  99  per  cent,  pure,  contains  but  1  to  2  per  cent,  of  impurities. 
That  this  commercial  cyanide  frequently  carries  a  varying  per- 
centage of  sodium  cyanide  is  a  well-known  fact;  and  it  of  course 
naturally  follows  that  the  greater  the  percentage  of  sodium  cya- 
nide contained  in  the  ordinary  98  to  99  per  cent,  potassium  cyanide, 
the  greater  the  percentage  of  impurities. 

As  soon  as  a  vat  is  charged,  from  20  to  25  tons  of  weak  solu- 
tion (carrying,  as  just  stated,  from  0.25  to  0.30  per  cent,  of  KCN) 
is  introduced,  from  the  bottom,  by  means  of  a  2-in.  drop-pipe, 
terminating  in  a  T  underneath  the  filter.  This  solution  is  in- 
troduced slowly  in  order  to  avoid  channeling  of  the  charge;  it 
usually  makes  its  appearance  on  top  of  the  sand  about  six  or 
seven  hours  after  being  turned  on.  When  the  solution  stands  2 
or  3  in.  above  the  top  of  the  charge,  it  is  turned  off,  and  the 
material  is  allowed  to  soak  for  six  hours,  during  which  time  the 
sand  will  usually  have  settled  from  3  to  4  in.  The  weak-solution 
discharge-valve  at  the  bottom  of  the  vat  is  now  opened  and 
leaching  is  commenced.  During  the  next  two  or  three  days 
weak  solution  is  added  from  the  top,  as  rapidly  as  permitted  by 
the  leaching  rate  of  the  charge,  until  a  total  of  from  100  to  130 
tons  has  been  applied.  From  60  to  70  tons  of  strong  solution, 
averaging  between  0.75  and  0.80  per  cent,  of  KCN,  is  now  run 
through  the  charge  at  a  somewhat  slower  rate,  the  usual  time 
consumed  by  this  operation  being  forty-eight  hours.  Weak  solu- 
tion is  next  run  through  the  charge  as  rapidly  as  possible,  until 
twenty-four  hours  before  the  time  it  is  to  be  discharged;  then 
wash-water,  to  the  amount  of  15  to  20  tons,  is  added  in  lots  of  5 
tons  each.  The  residue  is  then  ready  for  sluicing,  which  is  accom- 


296  HYDROMETALLURGY   OF  SILVER 

plished  by  two  men  in  about  six  hours,  each  using  a  2-in.  hose, 
equipped  with  a  0.5-in.  nozzle  and  operating  under  a  head  of  72 
ft.  After  finishing  the  sluicing,  the  canvas  filter  is  usually  swept 
clean  with  a  broom;  if  this  is  not  done,  it  is  found  that  the  filter 
cloth  clogs  with  fine  slime  and  the  rate  of  filtration  is  lowered. 
Each  vat  is  equipped  with  two  10  x  10-in.  square  bottom- 
discharge  doors. 

The  quantity  of  wash-water  used  is  regulated  by  the  balance 
of  the  solutions  on  hand.  Although  a  separate  zinc-box  is  pro- 
vided for  waste  solution,  it  is  seldom  used  except  during  the  rainy 
season;  during  other  parts  of  the  year  but  little  solution  is  run 
to  waste.  Only  two  of  the  leaching-vats  are  under  cover,  and 
during  the  rainy  season  it  becomes  necessary  to  run  a  certain 
percentage  of  the  solution  to  waste;  each  heavy  rain  gives  the 
exposed  vats  a  very  appreciable  quantity  of  water. 

It  has  been  my  experience  that  thorough  oxygenation  of  the 
material  is  a  very  desirable  feature  in  the  cyanidation  of  gold 
ores;  in  the  case  of  the  Palmarejo  ores,  this  is  essential  to  obtain 
the  best  results.  Due  to  the  fact  that  the  major  portion  of  the 
value  is  in  silver,  the  actual  weight  of  fine  metal  to  be  acted  upon 
is  much  greater  than  is  ordinarily  the  case  with  gold  ores. 

In  order  to  permit  as  much  air  as  possible  to  be  supplied  to 
the  sand  during  treatment,  the  solution  is  frequently  allowed  to 
drain  down  several  inches  beneath  the  surface  of  the  charge;  air 
is  thus  allowed  to  penetrate  the  material  to  this  depth.  It  is  our 
custom  to  assay  each  charge  every  twenty-four  hours,  after  the 
first  five  days  of  treatment.  Before  each  sampling,  the  solution  is 
allowed  to  drain  down  several  inches  below  the  surface  sand, 
thus  allowing  additional  opportunity  for  the  entrance  of  air  into 
the  upper  layer  of  the  charge. 

Under  the  most  favorable  conditions,  however,  the  air  drawn 
into  the  top  layer  can  have  but  little  effect  on  the  lower  half. 
It  is  doubtless  due  to  this  difference  in  aeration  of  the  upper  and 
lower  portions  that  the  lower  half  of  the  tailing  will  run  from  1 
to  2  oz.  of  silver  higher  than  the  upper  half.  Frequently  this 
difference  is  even  more  marked,  a  variation  of  3  or  4  oz.  being 
obtained  between  the  upper  foot  and  the  bottom  foot  of  the 
residue. 

To  overcome  this,  and  after  many  experiments,  some  time  ago 
the  practice  was  adopted  of  transferring  as  many  charges  as  pos- 


CYANIDATION   OF  AURIFEROUS  SILVER  ORES  297 

sible  from  one  vat  to  another  during  the  treatment.  To  transfer 
a  vat  means  the  loss  of  practically  twenty-four  hours  of  its  avail- 
able leaching  time,  because  it  is  necessary  to  drain  the  charge 
for  twelve  hours  before  commencing  to  transfer  it.  Also,  it  is 
necessary  that  one  of  the  adjoining  vats  be  empty  at  the  proper 
time  to  receive  the  transferred  charge.  By  careful  manipulation, 
at  present  about  one-third  of  the  total  number  of  charges  treated 
are  transferred.  When  the  two  additional  vats,  now  in  course  of 
erection,  shall  be  completed,  a  greater  number  of  charges  can  be 
transferred  and  the  additional  capacity  afforded  will  also  permit 
a  longer  treatment  to  offset  the  time  lost  in  transferral.  The 
transferring  is  done  by  contract  for  $16  a  vat,  which  is  equal  to 
16c.  per  ton.  While  being  transferred,  the  material  is  of  course 
given  a  thorough  exposure  to  the  air;  any  existing  lumps  are 
broken  up  by  the  shoveling;  and,  roughly  speaking,  the  bottom 
layer  of  the  original  charge  becomes  the  top  layer  of  the  trans- 
ferred charge.  Operations  are  usually  so  timed  that  the  trans- 
fer takes  place  while  the  strong  solution  is  in  contact  with  the 
material.  During  the  transfer,  100  Ib.  of  slaked  lime  is  evenly 
distributed  near  the  bottom  of  the  vat  receiving  the  transferred 
charge. 

The  charge  is  sampled  just  before  and  just  after  transferral, 
the  latter  sample  being  1  oz.  higher  in  silver  than  the  former,  a 
result  doubtless  due  to  the  fact  that  the  tendency  is  to  obtain  a 
larger  percentage  of  the  top  half  of  the  charge  than  of  the  lower 
half;  and,  as  heretofore  mentioned,  the  lower  half  of  the  original 
charge,  after  the  transferral,  becomes  practically  the  upper  half 
of  the  transferred  charge. 

The  first  solution  added  after  the  transfer  is  introduced  slowly 
from  the  bottom,  after  which  the  regular  routine  treatment  is 
continued.  The  value  of  the  effluent  solution  from  a  charge  is 
found  to  increase  immediately  after  the  charge  has  been  trans- 
ferred, such  increase  being  usually  from  2  to  3  oz.  of  silver  per 
ton  of  solution. 

In  general,  all  head-  and  tailing-samples  of  the  sand  are 
taken  with  a  1.5-in.  auger  at  12  to  18  different  points. 

The  record  of  a  single  charge,  which  is  representative  of  what 
is  regularly  obtained  in  ordinary  operations,  is  as  follows: 

Extraction,  96.27  per  cent,  of  gold  and  53.21  per  cent,  of 
silver. 


298  HYDROMETALLURGY  OF  SILVER 

Total  time  of  treatment,  including  charging  and  discharging, 
11  days. 

Solution  added:  weak,  261;  strong,  64;  wash-water,  15;  total, 
340  tons. 

All  tailing-samples,  with  the  exception  of  the  discharged  tail- 
ing, were  washed  before  assaying. 

During  transferral  a  sample  taken  from  upper  18  in.  of  charge 
assayed:  $0.50  of  gold,  9.20  oz.  of  silver;  and  one  from  the  lower 
18  in.  of  charge  assayed  $0.82  of  gold  and  11.92  oz.  of  silver. 

The  effluent  solution  from  the  leaching-vats  is  carried  to  the 
sump-tanks  by  two  separate  lines,  one  for  the  weak,  the  other  for 
the  strong  solution.  These  tanks  are  of  masonry  and  are  three  in 
number.  Two  of  them  (having  a  combined  capacity  of  65  tons) 
are  connected  and  serve  as  a  weak-solution  sump;  the  other, 
having  a  capacity  of  25  tons,  is  used  for  the  strong  solution. 
All  solution  draining  from  the  leaching-vats  is  passed  through  the 
zinc-boxes  before  being  returned  to  the  vats. 

The  proper  tonnage  of  strong  solution  is  maintained  by  deter- 
mining the  strength  of  the  effluent  solution  from  the  leaching- 
vats;  when  this  strength  reaches  0.35  per  cent,  of  KCN,  the 
solution  is  turned  into  the  strong-solution  sump.  As  a  working 
guide  for  maintaining  the  proper  alkalinity  of  stock  solutions, 
they  are  titrated  every  day  with  the  addition  of  about  5  c.c.  of 
strong  lime-water;  10  c.c.  of  cyanide  solution  is  used  in  all  titra- 
tions.  If  the  addition  of  lime-water  causes  a  difference  of  more 
than  0.5  Ib.  in  the  indicated  strength  of  the  solution,  the  quantity 
of  lime  added  to  the  sand  charged  into  the  leaching-vats  is 
increased. 

From  the  sump  the  solution  is  elevated  by  a  3-in.  centrifugal 
pump  (900  r.p.m.),  to  three  storage-tanks  at  the  head  of 
the  zinc-boxes,  a  vertical  distance  of  29  ft.  and  a  horizontal  dis- 
tance of  150  ft.  These  tanks  are  each  10  ft.  in  diameter,  8  ft. 
deep,  and  have  a  capacity  of  19  tons.  Two  of  these  tanks  are 
used  for  the  weak,  and  one  for  the  strong  solution.  The  solution 
from  the  vats  now  passes  through  the  zinc-boxes,  from  which  it  is 
led  to  three  storage-solution  tanks  beneath  the  boxes.  These 
storage-tanks  are  made  of  No.  9  sheet  steel,  each  being  15  ft.  in 
diameter,  6  ft.  deep,  and  with  a  capacity  of  33  tons.  Two  of 
them  are  used  as  strong-solution  storage-tanks;  the  other,  as  a 
v  eak-solution  storage-tank.  The  strong  solution  is  brought  to 


CYANIDATION  OF  AURIFEROUS  SILVER  ORES  299 

the  required  strength  by  adding  cyanide  to  the  last  compartment 
of  the  strong-solution  zinc-box,  which  is  reserved  for  this  pur- 
pose. No  cyanide  is  added  directly  to  the  weak  solution. 

Precipitation  of  Silver  and  Gold.  —  There  are  six  zinc-boxes, 
five  for  the  weak  and  one  for  the  strong  solution.  The  five  weak- 
solution  boxes  are  constructed  of  No.  10  sheet-steel,  and  are  2  ft. 
wide  and  18  ft.  long  over  all.  Each  box  contains  eight  compart- 
ments, each  compartment  having  an  available  zinc  capacity  of 
24  x  24  x  18  in.,  equivalent  to  6  cu.  ft.  Six  compartments 
only  are  filled  with  zinc  shavings;  therefore,  each  box,  when 
freshly  dressed,  contains  36  cu.  ft.  of  zinc  shavings,  making  a 
total  of  180  cu.  ft.  of  zinc  shavings  in  the  five  weak-solution 
boxes. 

The  strong-solution  zinc-box  consists  of  seven  individual 
round  boxes  or  compartments,  placed  in  series,  each  compart- 
ment being  28  in.  in  diameter  and  24  in.  in  depth,  and  having  an 
available  zinc  capacity  of  5  cu.  ft.  Only  six  of  the  compartments 
are  filled  with  shavings,  the  last  compartment  being  reserved  for 
the  addition  of  the  quantity  of  cyanide  required  to  bring  up  the 
strong  solution  to  standard  strength.  The  strong-solution  zinc- 
box  has,  therefore,  a  total  of  30  cu.  ft.  of  zinc  shavings. 

Records  are  kept  of  the  quantities  of  weak  and  strong  solu- 
tion daily  passing  through  the  boxes,  together  with  their  assay 
values  before  and  after  precipitation.  These  records  for  the  year 
(1904)  show  that  91,793  tons  of  weak,  and  22,251  tons  of  strong, 
solution  passed  through  the  boxes ;  this  is  equivalent  to  an  aver- 
age of  251  tons  of  weak,  and  61  tons  of  strong  solution  every 
twenty-four  hours.  During  this  period  the  flow  of  solution 
through  the  boxes  was  interrupted  on  various  occasions  for  a 
short  time,  due  to  the  ordinary  clean-ups,  dressing  of  the  boxes 
and  unavoidable  delays.  Without  taking  such  stoppages  into 
account,  the  average  rate  of  flow  through  the  boxes  equaled  1.4 
tons  of  weak  solution  per  twenty-four  hours  per  cubic  foot  of  shav- 
ings, and  2.03  tons  of  strong  solution  per  twenty-four  hours  per 
cubic  foot  of  shavings. 

The  actual  rate  of  flow  exceeds  these  figures,  as  it  is  assumed 
that  the  boxes  were  at  all  times  kept  dressed  with  the  maximum 
amount  of  shavings,  which  was  seldom  the  case. 

The  shavings  are  cut  on  an  ordinary  zinc  lathe,  from  No.  9 
sheet  zinc,  the  size  of  the  sheets  being  18  x  84  in.  Ordinarily, 


300  HYDROMETALLURGY  OF  SILVER 

six  sheets  are  wound  on  the  mandrel  of  the  lathe  for  one  cutting. 
One  boy,  working  twelve  hours,  cuts  sufficient  shavings  to  supply 
both  the  leaching  and  agitation  plants,  which  together  require  an 
average  of  120  Ib.  per  twenty-four  hours.  It  is  found  best  to 
keep  only  a  few  days'  shavings  on  hand;  freshly  cut  shavings  give 
better  results  than  those  which  have  been  cut  for  some  time. 
The  customary  practice  of  moving  the  zinc  from  the  lower  to  the 
upper  compartments,  when  dressing  the  boxes,  is  not  followed, 
fresh  zinc  being  added  as  required  to  the  top  of  each  compart- 
ment. 

The  strength  of  the  solution  running  through  the  weak  boxes 
will  average  between  0.25  and  0.30  per  cent,  of  KCN;  while  that 
of  the  solution  going  to  the  strong  zinc-box  will  average  between 
0.35  and  0.45  per  cent,  of  KCN. 

The  average  assay  values  per  ton  of  the  solutions  entering 
the  zinc-boxes  are  approximately  as  follows: 

Weak  solution,  $1  of  gold  and  2.25  oz.  of  silver;  strong  solu- 
tion, $1.24  of  gold  and  3.5  oz.  of  silver. 

It  is  seldom  that  any  trouble  is  experienced  with  the  precipita- 
tion of  the  contained  values.  As  a  rule,  the  precipitation  of  the 
gold  is  practically  perfect;  that  of  the  silver  averages  95  per  cent. 
When  precipitation  falls  off,  it  is  usually  due  to  the  presence  of 
an  accumulated  excess  of  lime  in  the  solution. 

Clean-up  of  Zinc-boxes.  —  On  account  of  structural  difficulties, 
it  is  necessary  to  handle  the  precipitates  more  than  is  desirable. 
The  boxes  are  cleaned  twice  a  month.  Before  commencing  on 
any  box,  clear  water  is  passed  through  it  a  sufficient  length  of 
time  to  displace  most  of  the  cyanide  solution;  this  requires  10  or 
15  minutes.  The  shavings  in  the  first  compartment  are  thor- 
oughly washed,  after  which  they  are  removed  and  the  water  bailed 
out  into  the  next  compartment.  The  precipitates  are  now  con- 
veyed by  buckets  to  the  clean-up  box,  where  they  are  passed 
through  a  20-mesh  screen.  A  small  percentage  of  "short"  zinc 
passes  through  this  screen,  but  the  greater  part  of  such  product 
is  here  separated  from  the  finer  precipitate  and  is  returned  to  the 
boxes.  The  first  compartment  is  now  filled  with  water;  the  zinc 
contained  in  the  other  compartments  is  gradually  transferred  to 
it  and  thoroughly  washed,  the  precipitates  from  each  compart- 
ment being  carried  to  the  clean-up  box  as  before  mentioned. 
To  minimize  the  oxidizing  effect  resulting  from  exposure  of  the 


CYANIDATION   OF  AURIFEROUS  SILVER  ORES  301 

wet  zinc  to  the  atmosphere,  the  washed  shavings  are  at  once 
placed  in  the  highest  vacant  compartment  of  the  zinc-box  and 
covered  with  solution. 

The  precipitate  accumulating  in  the  first  compartment  from 
the  washing  of  the  shavings,  after  settling  for  a  short  time,  is 
also  removed  to  the  clean-up  box.  This  latter  is  provided  with 
three  smaller  settling-boxes,  placed  in  series,  which  take  the 
overflow  from  it.  The  bottom  of  the  clean-up  box  is  tapped  by 
a  4-in.  drop-pipe,  which  discharges  directly  into  two  large  drying- 
pans  beneath. 

The  product  is  now  dried  as  much  as  is  practicable,  and  then 
mixed,  carefully  sampled,  assayed  and  sold  on  the  premises  to 
one  of  the  large  ore-buying  companies.  The  moisture  in  the 
dried  precipitate  has  averaged  0.27  per  cent,  during  the  past  year. 

The  clean-ups  are  bulky;  the  net  dry-weight  of  precipitate  in 
each  clean-up  averaged  between  1100  and  1200  Ib.  avoirdupois 
during  the  past  year. 

Considering  the  fact  that  the  precipitates  receive  no  treatment 
whatever  beyond  being  passed  through  a  20-mesh  screen,  and  the 
simple  drying,  as  above  mentioned,  it  is  rather  surprising  that 
they  carry  such  a  high  percentage  of  fine  metal.  During  1904 
the  assay  returns,  on  which  the  sale  of  the  precipitates  is  based, 
have  averaged  slightly  over  20,000  oz.  silver,  and  approximately 
$8000  of  gold,  per  short  ton.  By  actual  weight,  therefore,  the 
percentage  of  fine  metal  contained  in  the  dried  product  recovered 
throughout  this  period  was  approximately  68.57  percent,  of  silver 
and  1.33  per  cent,  of  gold,  making  a  total  of  69.90  per  cent,  of  both 
metals. 

Two  clean-ups  during  1904,  of  a  combined  net  weight  slightly 
exceeding  2300  Ib.,  gave  an  average  assay  value  of  22,200  oz. 
of  silver  per  ton,  making  the  fine-silver  content  equal  to  76.12 
per  cent,  by  weight  of  the  precipitates. 

A  record  of  the  labor  employed  in  clean-ups  shows  that  four 
men  (one  American  and  three  native  helpers)  would  readily  re- 
move 1200  Ib.  (net  dry-weight)  of  precipitate  from  the  boxes, 
and  have  the  product  in  the  drying-pans  in  eight  hours.  Based 
on  the  average  assays  of  the  precipitate  for  the  year,  this  means 
that  in  thirty-two  man-hours  approximately  12,200  oz.  of  fine 
metal  would  he  handled,  being  equivalent  to  a  duty  of  371  oz. 
per  man-hour.  This  rather  high  duty  is  of  course  due  entirely 


302  HYDROMETALLURGY  OF  SILVER 

to  the  fact  that  the  precipitate  contains  high  percentages  of  pre- 
cious metals. 

Tonnage  and  Extraction.  —  During  1904,  34,900  tons  of  sand 
were  treated  in  the  leaching-plant.  This  tonnage  would  have 
been  considerably  greater  had  it  not  been  that  during  this  period, 
aside  from  the  stoppages  due  to  general  repairs,  the  mill  was  closed 
down  for  intervals  aggregating  57  days  for  the  entire  50  stamps. 

The  extraction  for  1904  (as  indicated  by  the  assay  differences 
between  the  sand  charged  into  the  leaching-vats  and  that  being 
discharged)  was  95.5  per  cent,  of  the  gold  and  52.5  per  cent,  of  the 
silver  value.  The  combined  total  during  the  year  checks  closely 
with  that  called  for  by  the  sand  assays.  The  assay  value  of  the 
sand  treated  during  this  period  averaged  $2.85  of  gold,  and 
slightly  more  than  16  oz.  of  silver  per  ton.  During  1904,  the  re- 
turns from  the  precipitates  were  practically  1  per  cent,  less  in  gold 
and  0.5  per  cent,  more  in  silver  than  those  called  for  by  the  pre- 
cipitation records. 

Consumption  of  Cyanide,  Zinc  and  Lime.  —  The  office  records 
show  that  for  1904  the  consumption  per  ton  of  sand  cyanided  was 
as  follows:  cyanide,  2.95  lb.;  zinc,  0.96  lb.;  and  lime,  4.33  Ib. 
Expressed  in  terms  of  potassium  cyanide,  this  consumption 
would  equal  3.69  lb.  of  potassium  cyanide  per  ton  of  ore  treated. 

General  Remarks.  —  The  total  quantity  of  solution  passing 
through  the  zinc-boxes  during  1904,  divided  by  the  quantity  of 
sand  for  the  same  period,  shows  that,  for  each  ton  treated,  3.27 
ton  of  solution  left  the  leaching-vats,  of  which  2.63  tons  was  weak 
and  0.64  ton  strong  solution.  It  is  found  that  a  large  quantity 
of  weaker  solution  gives  more  satisfactory  results  than  a  small 
quantity  of  the  strong,  and  it  is  always  made  an  important  point 
to  pass  as  much  weak  solution  through  a  charge  as  possible. 
Experience  has  demonstrated  that,  in  a  given  length  of  treat- 
ment, a  rapid  leaching  rate  and  a  large  quantity  of  solution  are 
more  efficient  than  a  slower  leaching  rate  and  a  consequently 
lesser  quantity  of  solution.  The  solution  pipe-lines  and  launders 
occasionally  become  quite  choked  in  places  with  scale  deposited 
from  the  solution.  This  scale,  taken  from  lines  carrying  precipi- 
tated solution,  contained  from  a  trace  to  $1  of  gold  and  from  1 
to  7  or  8  oz.  of  silver  per  ton.  The  scale  deposited  from  the  un- 
precipitated  solution  usually  runs  higher,  several  assays  taken 
having  averaged  about  $5  of  gold  and  18  oz.  of  silver  per  ton. 


CYANIDATION   OF  AURIFEROUS  SILVER  ORES          303 

Ordinarily  the  solutions  do  not  become  excessively  fouled. 
They  contain  small  quantities  of  iron  and  manganese  in  addition 
to  the  zinc  compounds  present.  Alkaline  sulphides  are  very 
rarely  or  never  noticed  in  solution;  however,  sulphocyanates  and 
ferrocyanides  appear  to  be  constantly  present  in  fair  quantities, 
about  0.41  per  cent,  of  ferrocyanide  and  0.048  per  cent,  of  sulpho- 
cyanate. 

The  sand  charged  averaged  about  0.09  per  cent,  of  latent 
acidity;  and,  as  a  rule,  it  contained  no  free  acid. 

The  concentrate  produced  is  sold  to  the  same  company  that 
buys  the  cyanide  precipitate.  An  attempt  had  been  made  to 
treat  the  concentrate  by  cyanide,  but  without  success.  Experi- 
ments on  both  raw  and  dead-roasted  concentrate  (reduced  to 
various  degrees  of  fineness)  by  leaching  and  agitation,  for  vary- 
ing periods  of  time  up  to  34  days,  and  using  solutions  varying 
from  0.2  to  2  per  cent,  of  KCN,  proved  unsatisfactory. 

Table  I,  given  herewith,  shows  the  working  costs  for  milling 
and  cyaniding  during  1904.  The  cost  of  all  supplies  is  increased 
by  the  heavy  freight  transportation  expenses,  as  well  as  by  the 
duties  placed  by  the  Mexican  Government  on  most  of  the  supplies 
used. 

TABLE  I  — WORKING  COSTS  PER  TON 
Milling: 

Supplies $0.640 

Labor 0.357 

Lubricating 0.023 

Assay  office  (labor  and  supplies) • 0.035 

Concentrating 7~: .  .  .  .     0.092 

Power  (ditch  maintenance  and  supplies)   0.234 

Salaries    0.264 

Miscellaneous  (lighting,  etc.)  0.018 

Management  and  general  expenses   0.336 

Total  *   $1.999 

Cyaniding: 

Cyanide  (2.95  Ib.  @  $0.63)  $1.859 

Zinc  (0.96  Ib.  @  $0.30) 0.288 

Lime  (4.33  Ib.  @  $0.0118)    0.051 

Other  supplies 0.050 

Labor  .  ...  0.329 

Salaries    0.371 

Assay  office  (labor  and  supplies) 0.036 

Power  (ditch  maintenance  and  supplies) 0.017 

Miscellaneous  (lighting,  etc.)    0.004 

Management  and  general  expenses 0.186 

Total  2   $3.191 

1  $1.999  Mexican  currency  during  this  period  was  equivalent  to  $0.95  gold. 

2  $3.191  Mexican  currency  during  this  period  was  equivalent  to  $1.22  gold. 


304  HYDROMETALLURGY  OF  SILVER 

The  cost  of  realization  on  cyanide  precipitate  has  not  been 
included  in  the  cyanide  working  costs  given  herewith.  Trans- 
portation expenses  on  the  precipitates  are  also  very  heavy. 
In  addition  to  these  comes  the  heavy  item  of  government  bullion 
taxes.  The  average  cost  of  realization  on  cyanide  precipitate, 
per  ton  of  ore  cyanided,  is  as  follows:  Government  taxes,  $0.84; 
treatment  charges  (including  transportation  expenses),  $1.06; 
total,  $1.90.  The  cost  of  realization  on  the  concentrate  produced 
is  also  unusually  high,  on  account  of  the  heavy  transportation  ex- 
penses and  government  bullion  taxes.  The  average  cost  of 
realization  per  ton  of  ore  crushed  is  as  follows:  Government  taxes, 
$0.35;  treatment  charges  (including  transportation  expenses), 
$1.08;  total,  $1.43. 

Treatment  of  Slime.  —  As  already  mentioned,  the  accumulated 
and  currently  produced  slime  is  now  being  treated  in  a  separate 
plant  by  a  system  of  agitation  and  decantation,  centrifugal 
pumps  being  used  as  the  means  of  agitation.  The  slime  plant 
consists  essentially  of  the  following  parts  and  accessories: 

Four  agitation  and  four  decantation  vats,  each  provided  with 
conical  bottoms  and  connected  with  its  own  separate  centrifugal 
pump;  two  solution  tanks  placed  at  the  head  of  the  zinc-boxes, 
which  receive  the  solution  from  the  decantation  vats;  four  sets 
of  zinc-boxes  and  three  solution  sumps,  which  receive  the  solu- 
tion leaving  the  zinc-boxes;  one  special  solution  tank,  placed 
at  a  higher  level  than  the  rest  of  the  plant,  and  used  prin- 
cipally to  supply  solution  to  the  pump  bearings  under  pressure; 
two  ordinary  3-in.  centrifugal  pumps,  used  only  for  pumping 
solution  from  the  sumps  to  any  desired  vat  or  to  the  upper 
solution  tank  just  mentioned,  they  being  so  connected  up  that 
either  pump  can  be  used  should  the  other  get  out  of  order. 
Each  pump  is  run  by  a  friction-clutch  pulley,  which  enables 
it  to  be  started  or  stopped  in  a  moment,  independently  of  the 
other  pumps.  A  small  14  x  15-in.  friction-geared  hoist  is  used 
to  convey  the  slime  from  the  slime  pits  to  the  agitation  vats. 
The  entire  plant  is  run  by  a  5-ft.  Pelton  wheel,  making  about 
115  r.p.m.,  and  operating  under  a  head  of  81  ft.,  using  a  4-in. 
nozzle.  Water-power  is  obtained  by  means  of  a  14-in.  riveted 
steel  pipe,  tapping  the  main  pipe  line  supplying  power  to  the 
mill.  This  14-in.  pipe-line  was  brought  in  by  mule  back,  riveted 
in  10-ft.  lengths,  although  some  difficulty  was  experienced  in  its 


CYANIDATION  OF  AURIFEROUS  SILVER  ORES 


305 


transportation.     (Figs.  76  and  77  give  plan  and  section  of  the 
slime  plant.) 

The  method  of  treating  the  slime  is  similar  to  that  ordinarily 
practised  by  agitation  and  decantation;  it  consists  in  giving  a 
two  days'  agitation  in  the  agitation  vats,  with  from  two  to  three 
times  the  weight  of  cyanide  solution,  followed  by  another  two 
(or  sometimes  three)  days'  treatment  in  the  decantation  vats; 
during  the  latter  treatment,  the  charge,  after  having  been  suffi- 
ciently agitated  with  the  addition  of  slaked  lime,  is  allowed  to 
settle  as  much  as  practicable,  and  the  clear  supernatant  liquor  is 
decanted  and  passed  through  the  zinc-boxes.  This  operation 
of  agitation,  settling  and  decantation  of  clear  solution  is  repeated 
as  many  times  as  permissible  within  the  time  limit  of  the  treat- 
ment, being  ordinarily  three  or  four  decantations. 

The  material  treated,  when  dried  to  20  per  cent,  or  25  per  cent, 
moisture,  is  tough  and  of  the  consistency  of  soft  putty.  It  con- 
tains, however,  a  certain  percentage  of  fine  sand  and,  when  viewed 
in  vertical  section,  presents  a  somewhat  stratified  appearance. 
On  long  drying,  it  cracks  into  layers  which  are  almost  absolutely 
impervious  to  leaching. 

The  results  of  the  sizing-test  (which  are  given  in  Table  II) 
represent  an  average  obtained  from  the  material  treated. 

TABLE  II  — SIZING-TEST  ON  SLIME 

Assay  value  of  material  was  $4.13  of  gold  and  20.30  oz.  of  silver  per  ton. 


SIZE  OF  MATERIAL 

WEIGHT 

ASSAY  VALUE 

PERCENTAGE  OF 
TOTAL  VALUES 
CONTAINED 

GOLD 

SILVER 

GOLD 

SILVER 

Retained  on   80  mesh.    .  . 
100           ... 
120 
150 
200 
Passed           200 

Per  Ct. 
1.1 
2.7 
5.6 
3.1 
2.7 
84.8. 

$2.38 
2.06 
1.96 
2.27 
2.16 
4.54 

Oz. 
14.22 
13.60 
13.02 
14.14 
13.10 
21.68 

Per  Ct. 
0.83 
1.35 
2.66 
1.70 
1.41 
93.22 

Per  Ct. 
0.77 
1.81 
3.59 
2.16 
1.74 
90.52 

Totals  

100.0 

100.92 

100.59 

Description  of  the  Plant.  —  The  four  agitation  vats,  constructed 
of  3-in.  redwood  throughout,  are  provided  with  conical  bottoms, 
slanting  at  45  deg.  As  shown  in  Fig.  78,  each  vat  has  an  inside 
diameter  of  15  ft.  7  in.;  the  vertical  depth,  from  top  of  side  staves 


306 


HYDROMETALLURGY  OF  SILVER 


FIG.  76.  — PLAN  OF  SLIME   PLANT. 


CYANIDATION  OF  AURIFEROUS  SILVER  ORES  307 


308 


HYDROMETALLURGY   OF  SILVER 


to  the  iron  casting  at  point  of  conical  bottom,  is  15  ft.,  the  inside 
depth  of  vertical  side  staves  being  7  ft.  3  in.  Each  agitation  vat 
is  connected  with  a  special  manganese-steel  lined  4-in.  centrifugal 
pump,  which  runs  at  a  speed  of  900  r.p.m.  The  pump  is  con- 
nected with  the  vat  by  the  4-in.  suction  pipe  a,  which  enters  the 
vat  through  the  side  staves  about  6  in.  above  their  juncture  with 
the  bottom  staves  and  extends  nearly  to  the  center  of  the  vat, 
where  it  is  connected  by  means  of  a  movable  elbow,  6,  with  a 
short  piece  of  4-in.  pipe,  c,  provided  at  the  free  end  with  a  good- 
sized  screen  or  strainer,  d,  made  of  J-in.  sheet  iron,  punched  with 
a  number  of  1-in.  holes;  this  short  piece  of  pipe,  together  with 


FIG.  78.  — AGITATION  VAT. 

the  screen,  is  of  such  a  length  that  when  being  lowered  the  screen 
will  just  clear  the  bottom  staves.  The  screen  is  provided  with  a 
small  iron  ring,  to  which  is  fastened  a  piece  of  rope,  by  means 
of  which  it  can  be  raised  and  lowered. 

Just  outside  the  vat,  the  suction-pipe  is  provided  with  an 
air-cock,  e,  which  admits  air  to  the  material  going  through  the 
pump.  This  air-cock,  however,  is  rarely  used  at  the  present 
time.  The  service-cock  /  permits  the  shutting  off  of  the  material 
from  the  pump  at  any  time  it  may  become  necessary,  as,  for  in- 
stance, to  repack  the  stuffing-box  or  to  examine  the  interior  of 


CYANIDATION  OF  AURIFEROUS  SILVER  ORES  309 

the  pump.     The  2-in  pipe-line  g,  provided  with  the  valve  h,  con- 
nects with  the  upper  solution  tank. 

When  it  becomes  necessary  to  shut  the  pump  down  for  any 
length  of  time  (either  at  the  conclusion  of  the  agitation  of  the 
charge  or  at  any  time  during  the  treatment),  the  2-in.  valve  h 
is  opened  and  the  clear  solution  only  passes  through  the  pump. 
The  friction-clutch  pulley  running  the  pump  is  now  thrown  out 
of  clutch,  and  after  the  pump  has  stopped  the  valve  h  is  closed. 
By  this  means  is  avoided  the  accumulation  in  the  pump  interior 
of  solid  matter  that  would  naturally  be  deposited  (when  the 
pump  is  stopped  for  any  length  of  time)  from  the  slimy  material 
ordinarily  passing  through  it. 

The  4-in.  discharge-pipe  i  of  the  pump  is  provided  with  a  small 
bib-nosed  petcock,  /,  a  few  inches  from  the  body  of  the  pump, 
by  means  of  which  samples  can  readily  be  taken  of  the  material 
passing  through  the  pump.  The  discharge-pipe  passes  over  the 
top  of  the  vat,  and  at  a  point  vertically  over  the  center  of  the 
bottom  casting  is  provided  with  an  elbow  and  drop-pipe,  k,  which 
reaches  to  within  about  15  in.  of  the  bottom  casting.  This  pipe 
is  held  firmly  in  position  by  means  of  an  iron  clamp  and  four  legs 
made  of  f-in.  bolts  fastened  to  the  bottom  casting  and  which 
serve  as  a  support.  The  distance  of  the  lower  end  of  this  dis- 
charge-pipe from  the  bottom  of  vat  is  a  matter  of  some  impor- 
tance in  the  agitation,  and  a  number  of  experiments  made  along 
this  line  have  indicated  that  the  best  satisfaction  is  obtained  at  a 
distance  of  15  in.  from  the  bottom  casting. 

Different  shapes  of  nozzles  have  been  tried  at  the  lower  end 
of  the  drop-pipe,  but  experience  has  shown  that  the  plain  4-in. 
pipe-end  gives  satisfactory  results.  The  discharge-pipe  of  the 
pump  tends  to  act  as  a  siphon  when  the  pump  is  stopped  at  any 
time  during  the  agitation,  and  would  therefore  cause  inconvenience 
when  repacking  the  stuffing-box  or  making  any  necessary  repairs. 
To  prevent  this,  air  is  admitted  to  the  pipe  by  opening  the  small 
air-cock  I,  tapped  into  the  elbow  at  the  upper  end  of  the  drop- 
pipe.  This  air-cock  I  is  also  frequently  used  to  allow  the  air 
entering  into  the  charge  to  be  agitated,  it  being  found  for  this 
purpose  preferable  to  the  air-cock  on  the  suction-pipe.  It  might 
be  supposed  that  when  this  air-cock  is  open  during  the  agitation  a 
steady  stream  of  the  material  passing  through  the  discharge-pipe 
would  be  ejected  through  it;  and  with  regard  to  the  air-cocks 


310  HYDROMETALLURGY  OF  SILVER 

similarly  situated  on  the  pump  connections  of  the  decantation 
vats,  such  is  the  case.  As  regards  the  pumps  connected  with 
the  agitation  vats,  however,  the  effect  is  found  to  be  quite  the 
reverse,  and  a  rather  strong  air  suction  usually  occurs  when  this 
air-cock  is  open. 

The  pump-bearing  nearest  the  pump  shell  is  tapped  with  a 
small  pipe-line,  m,  provided  with  the  valve  n,  which  connects  with 
the  upper  solution  tank.  By  this  means  the  bearing  is  supplied 
with  clear  solution  under  pressure,  and  the  wear  on  the  shaft  and 
bearing  is  greatly  reduced.  At  the  commencement  of  operations 
clear  water  was  supplied  to  the  pump  bearings  and  was  also  used 
for  cleaning  out  the  pumps  and  for  priming,  when  necessary. 
It  was  soon  found,  however,  that  the  quantity  of  water  added  in 
this  way  increased  the  volume  of  stock  solution  very  appreciably, 
and,  of  course,  an  equal  quantity  of  weak  cyanide  solution  had 
ultimately  to  run  to  waste.  Not  only  did  this  cause  an  unneces- 
sary mechanical  consumption  of  cyanide,  but  the  quantity  of 
water  added  through  the  pump  bearings  naturally  reduced  the 
strength  of  the  working  solution  in  the  vat  under  operation,  with 
a  deleterious  effect  on  the  percentage  of  extraction.  The  quantity 
of  solution  that  will  be  added  to  a  vat  during  the  usual  period 
of  agitation  (from  forty  to  forty-four  hours),  when  the  shaft  and 
bearing  are  a  little  worn,  is  surprising,  amounting  in  some  cases  to 
15  tons,  even  when  the  greatest  care  is  exercised.  The  amount  of 
solution  added  in  this  way  is  naturally  the  least  just  after  the 
pump  has  been  equipped  with  a  new  shaft  and  new  liners,  and 
the  bearing  rebabbitted.  On  an  average,  however,  the  quantity 
of  solution  added  to  each  charge  through  the  pump  bearings  is 
from  5  to  6  tons.  The  agitation  pumps  in  use,  while  in  most 
respects  proving  very  satisfactory,  have  nevertheless  certain  de- 
fects in  their  design,  which  contribute  largely  to  the  rapid  wear- 
ing of  the  shaft  and  the  bearing  next  to  the  pump  shell,  and  also 
to  the  wearing  of  the  interior,  renewable  manganese-steel  wear- 
ing parts.  The  life  of  these  parts  naturally  varies;  but  ordinarily 
it  is  necessary  to  equip  a  pump  with  a  new  shaft  and  with  por- 
tions of  the  manganese-steel  wearing  parts,  and  to  rebabbit 
the  bearing  every  six  weeks.  The  pumps  are  equipped  with  a 
pulley  having  a  6-in.  face;  but  it  is  found  preferable  to  use  a 
4-in.  belt,  since  this  reduces  the  weight  on  the  pump-shaft  with 
a  consequent  decrease  in  its  wear,  while  a  4-in.  belt  runs  the 


CYANIDATION  OF  AURIFEROUS  SILVER  ORES 


311 


pump  equally  as  well  as  a  6-in.  one.     Wire  lacing  is  used  on  all 
the  belts. 

Each  agitation  vat  was  originally  provided  with  a  6-in.  dis- 
charge opening  at  the  center  of  the  bottom  casting.  This  open- 
ing was  bushed  down  to  4  in.  and  was  provided  with  a  nipple 
and  a  straightway  valve.  The  first  few  vats  were  discharged 
from  the  bottom  by  this  means,  but  a  deal  of  trouble  was  experi- 
enced, due  to  the  fact  that,  though  all  the  slime  entering  the 
agitation  vats  passed  through  a  grizzly  having  1.25-in.  openings, 
yet  the  bottom  valve  would  frequently  become  choked  with  small 
rocks  and  other  material  which  seemed  to  be  mixed  with  the 
first  slime.  This  bottom  discharge  was  therefore  discontinued, 


FIG.  79.  —  DECANTATION  VAT 

a  hole  was  bored  in  the  bottom  staves  10  in.  from  the  bottom 
casting,  and  a  3.5-in.  iron  service-cock  was  secured  to  the  vat  by 
means  of  a  short  nipple  and  iron  flanges.  The  vats  are  discharged 
through  this  valve  into  a  wooden  launder  which  conveys  the 
material  to  the  corresponding  decantation  vat.  This  launder  is 
provided  with  rows  of  6-in.  wire  nails,  which  serve  to  catch  any 
foreign  matter. 

The  four  decantation  vats,  made  of  3-in.  redwood  through- 
out, are  of  the  same  dimensions  as  the  agitation  vats,  with  the 
exception  that  they  are  provided  with  conical  bottoms,  slanting 
at  20  deg.  Each  one  is  connected  with  an  ordinary  3-in.  centrif- 
ugal pump.  Fig.  79  shows  in  detail  the  connection  of  the  pump 


312  HYDROMETALLURGY   OF  SILVER 

with  the  vat.  The  vat  is  discharged  through  a  3.5-in.  bottom- 
discharge  valve  and  pipe,  into  the  residue  launder,  from  which 
the  discharged  material  flows  to  the  river.  Removal  of  the  clear 
solution  is  effected  by  means  of  a  2-in.  decantation  pipe  and  float. 
This  pipe  enters  the  side  of  the  vat  about  6  in.  above  the  bottom 
staves  and  is  provided  writh  two  loosely  threaded  elbows,  which 
permit  of  the  free  raising  and  lowering  of  the  portion  within  the 
vat.  The  float  proper  is  made  of  two  ordinary  5-gal.  oil  cans, 
soldered  water-tight  and  painted  with  paraffin  paint.  The  rate 
of  decantation  is  controlled  by  means  of  the  2-in.  valve  just 
outside  of  the  vat. 

It  frequently  happens  that  the  solution  drawn  from  the  decan- 
tation vats  is  not  perfectly  clear,  and  two  filter-boxes  are  pro- 
vided (see  Figs.  76  and  77)  for  the  partial  clarification  of  the 
solution  before  it  enters  the  solution  tanks  at  the  head  of  the 
zinc-boxes.  Each  compartment  of  these  filter-boxes  is  provided 
with  a  discharge  valve,  by  means  of  which  the  sediment  deposited 
from  the  solution  can  be  washed  into  a  waste  launder. 

The  solution  tanks  at  the  head  of  the  zinc-boxes  are  two  in 
number,  one  being  used  for  the  weak  and  the  other  for  the  strong 
solution.  They  are  made  of  2-in.  redwood  throughout,  and  are 
each  11  ft.  8  in.  in  diameter  and  7  ft.  7  in.  deep,  inside  measure- 
ments, having  a  capacity  of  25  tons.  Each  solution  tank  is 
provided  with  a  2-in.  floating  hose,  by  means  of  which  the  clearest 
solution  in  the  tanks  is  always  supplied  to  the  zinc-boxes.  A  3-in. 
valved  opening  in  the  bottom  of  each  of  these  solution  tanks  per- 
mits of  the  discharge  of  the  accumulated  slime  into  a  waste 
launder. 

Fig.  80  shows  the  timber  foundations  supporting  the  decanta- 
tion vats,  the  conical  bottoms  resting  on  three  beveled  rings. 
The  supports  for  the  agitation  vats  are  built  in  the  same  manner, 
the  supporting  rings,  however,  being  placed  to  line  at  45  deg. 
instead  of  at  20  deg.  Fig.  81  shows  the  decantation  vats  in  course 
of  erection. 

.  There  are  four  sets  of  zinc-boxes,  each  set  being  composed  of 
six  round  individual  boxes  or  compartments,  each  compartment 
being  28  in.  in  diameter  and  2  ft.  deep,  and  having  an  available 
zinc  capacity  of  approximately  5  cu.  ft.  One  of  the  boxes  is 
used  solely  for  strong  solution,  and  two  for  weak  solution;  the 
fourth  being  so  connected  up  that  either  weak  or  strong  solution 


FIG.  80.  —  Timber  Foundations  supporting  Decantation  Vats  of  Slime  Plant. 


FIG.  81  —  Decantation  Vats  in  Course  of  Construction. 


CYANIDATION  OF  AURIFEROUS  SILVER  ORES  313 

may  be  run  through  it.  The  solution  leaving  the  zinc-boxes 
passes  to  three  sump-tanks,  made  of  2-in.  redwood  throughout, 
each  11  ft.  8  in.  in  diameter  and  9  ft.  7  in.  deep,  inside  measure- 
ments, and  of  a  capacity  of  32  tons  of  solution.  Two  of  these 
tanks  are  connected  together  and  serve  as  a  weak-solution  sump, 
the  other  being  used  for  the  strong  solution. 

Fig.  82  gives  a  good  view  of  the  plant  shortly  before  its  com- 
pletion and  shows  its  general  arrangement.  Fig.  83  gives  a 
nearer  view  of  three  of  the  agitation  vats  and  shows  the  tops  of 
two  of  the  decantation  vats.  The  4-in.  centrifugal  pump  (lead- 
ing to  No.  1  agitation  vat)  is  seen  partially  connected  up.  Pro- 
truding from  the  top  of  the  decantation  vat,  a  little  below  the 
center  of  the  picture,  is  seen  the  end  of  one  of  the  2-in.  decanta- 
tion pipes. 

Method  of  Treatment.  —  The  accumulated  slime,  after  having 
been  dried  in  the  slime  pits  as  much  as  practicable,  is  conveyed 
to  the  agitation  vats  in  ordinary  half-ton  ore  cars  by  means  of  a 
small,  friction-geared  hoist.  Each  agitation  vat  is  provided 
with  an  iron  grizzly  (measuring  3  ft.  3  in.  x  9  ft.,  and  having 
1.25-in.  openings),  which  is  suspended  over  to  one  side  of  the 
center.  The  content  of  the  car  is  dumped  on  to  this  grizzly  and 
the  portion  that  does  not  pass  of  its  own  weight  is  trampled,  or 
otherwise  forced  through,  by  boys.  For  some  time  the  material 
being  treated  averaged  from  20  to  25  per  cent,  of  moisture  and  in 
this  condition  was  lumpy  and  cohesive.  During  this  period  the 
agitation  was  unsatisfactory  and  the  percentage  of  extraction 
was  low.  Difficulty  was  experienced  in  discharging  the  vats; 
the  unagitated  portion  of  the  charge  would  remain  in  the  pointed 
bottom  of  the  vat  as  a  tough,  putty-like  mass,  after  all  the  liquid 
portion  had  been  discharged,  and  could  only  be  washed  out  by 
means  of  a  stream  of  solution  or  water  under  pressure.  Experi- 
ence demonstrates  that  the  best  condition  of  the  material  is  such 
that,  when  dumped  on  the  grizzly,  it  will  run  through  of  its  own 
weight.  In  this  state  the  slime  carries  from  30  to  35  per  cent, 
moisture.  It  is  desirable  that  the  percentage  of  moisture  con- 
tained in  the  slime  when  charged  shall  be  as  low  as  possible,  com- 
patible with  satisfactory  agitation;  the  greater  the  percentage  of 
moisture  contained  in  the  slime,  the  greater  will  be  the  mechanical 
consumption  of  cyanide.  The  complete  drying  of  the  slime  by 
some  cheap  process,  followed  by  powdering  before  charging  into 


314 


HYDROMETALLURGY   OF  SILVER 


the  agitation  vats,  should  be  productive  of  improved  results. 
A  charge  equivalent  to  about  15  tons  of  dry  slime  gives  more 
satisfactory  results  than  does  a  heavier  one. 

TABLE    III.  — SETTLING    RATE    OF    SLIME    PER    HOUR,    WITH 
ADDITION  OF  LIME 

SETTLEMENT  (IN  INCHES)  OF  SLIME 


Test 
No.  1 

Test 
No.  2 

Test 
No.  3 

Test 
No.  4 

Test 
No.  5 

(e) 

Test 
No.  6 

(<0 

Test 
No.  7 

(c) 

Proportion    of    solution    to 
slime 

2.5:1 

2.5:1 

2.5:1 

2.5:1 

2.5:1 

2.5:1 

3.3:1 

Lime    added    per    ton    of 
slime  (<z)  .               ... 

21b. 

31b. 

31b. 

31b. 

31b. 

None 

(6) 

41b. 

1  hour  

11.0 

10.5 

10.0 

16.0 

14.0 

15.0 

22.0 

2  hours 

21.0 

190 

165 

250 

21  0 

24  5 

365 

3 

27.5 

26.0 

23.5 

33.0 

300 

335 

51  5 

4 

33.0 

32.0 

30.0 

40.0 

390 

400 

54  0 

5           

36.0 

35.5 

36.0 

42.0 

43.0 

42.0 

570 

6           
7          

38.0 
39.5 

38.5 
40.0 

40.0 
41.5 

43.0 
44.0 

47.0 

48.5 

43.0 
44.0 

58.0 
590 

8          

40.5 

41.0 

42.5 

44.5 

48.5 

44.5 

59.5 

9 

41.0 

41  5 

430 

45  0 

49  0 

44  5 

59  5 

10 

41.0 

42  5 

43  5 

45  0 

49  0 

45  0 

59  5 

11           .... 

41.5 

42  5 

43.5 

450 

12           

41.5 

43.0 

(a)  This  quantity  of  lime  was  added  in  addition  to  the  lime  already  con- 
tained in  the  solution;  sufficient  lime  usually  being  present  in  solution  that 
the  addition  of  5  c.c.  of  strong  lime  water  to  a  titration  (with  silver  nitrate), 
for  strength  of  solution,  would  make  no  difference  in  the  titration. 

(6)  See  note  (a). 

(c)  Tests  No.  5  and  No.  6  were  on  material  from  near  the  head  of  slime 
pits,  and  which  therefore  contained  a  larger  percentage  than  usual  of  fine 
sand. 

(Each  2  in.  of  solution  equals  one  ton.) 

Before  commencing  to  charge  the  slime,  about  35  tons  of  solu- 
tion from  the  strong  solution  sump  (usually  of  a  strength  between 
0.12  and  0.15  of  KCN)  is  pumped  into  the  vat  and  the  attached 
centrifugal  pump  started.  From  75  to  100  Ib.  of  slaked  lime  is 
added  and  the  charging  of  the  slime  is  commenced.  After  the 
required  quantity  of  slime  has  been  added,  a  sample  of  the  ma- 
terial passing  through  the  pump  is  taken,  filtered  and  the  clear 
solution  titrated.  The  necessary  quantity  of  cyanide  to  bring 
the  solution  up  to  strength  is  then  added.  Experiments  have 
been  made  with  various  strengths  of  solution  in  the  agitation  vats; 


•  4 


fcl 


i: 

iU'*fc       • 


FIG.  82  —  General  Arrangement  of  Slime  Plant. 


FIG.  83  —  Three  of  the  Agitation  Vats  and  Tops  of  two  of  the  Decantation  Vats. 


CYANIDATION  OF  AURIFEROUS  SILVER  ORES  315 

the  results  thus  far  show  the  0.2  per  cent,  solution  to  give  more 
satisfactory  results  than  the  use  of  a  weaker  solution.  The  solid 
cyanide  is  placed  in  perforated  buckets  or  cans  and  suspended 
in  the  charge.  It  is  found  that  unless  the  receptacles  containing 
the  cyanide  be  frequently  agitated  about  in  the  charge,  the  cya- 
nide dissolves  exceedingly  slowly.  The  less  the  proportion  of 
solution  to  solid  matter  present,  the  more  noticeable  is  this  ten- 
dency of  the  cyanide  to  dissolve  slowly.  It  is  also  noticed  that, 
the  thicker  the  charge,  the  slower  is  the  action  of  the  cyanide  on 
the  silver  and  gold  contained  in  the  slime.  During  agitation  it  is 
best  to  keep  the  screen  at  the  end  of  the  suction  pipe  just  as  near 
the  surface  of  the  charge  as  possible,  without  allowing  the  entrance 
of  air.  By  so  doing,  the  material  passing  through  the  pump 
always  contains  a  minimum  quantity  of  solids,  and  the  wear  on 
the  pump  is  consequently  lessened.  In  addition  to  this,  the 
movement  or  circulation  within  the  charge  is  then  greatest,  since 
the  suction  and  discharge  points  are  then  most  separated.  It  is 
quite  probable  that  a  considerable  portion  of  the  heaviest  and 
coarsest  part  of  the  material  treated  does  not  pass  through  the 
pump  at  all;  as,  owing  to  its  greater  weight,  it  may  never  be 
raised  to  the  hight  of  the  suction  screen.  The  agitation  of  the 
mass  seems  to  depend  chiefly  on  the  fact  that  the  discharge  issu- 
ing from  the  drop-pipe  tends  to  keep  the  point  of  the  conical 
bottom  free  from  any  settled  deposit  of  slime,  and  the  thickened 
material,  constituting  the  lower  portion  of  the  charge,  keeps 
constantly  sliding  down  the  inclined  sides  toward  the  bottom 
point.  The  product  issuing  from  the  discharge-pipe,  being  drawn 
from  the  surface  of  the  charge,  must  pass  upward  through  the 
entire  mass  above,  before  it  can  again  pass  through  the  pump. 

The  percentage  of  solid  matter  contained  in  the  material 
passing  through  the  agitation  pumps  is  determined  from  samples 
taken  through  the  bib-nosed  petcock  tapping  the  discharge-pipe 
a  few  inches  above  the  pump  shell.  The  pulp  passing  through 
the  pumps  will  carry  25  per  cent.,  by  weight,  of  solids. 

A  thorough  oxygenation  of  the  mass  is  found  to  be  an  essential 
feature;  it  becomes  more  necessary  as  the  proportion  of  solid 
matter  to  solution  present  increases.  At  the  commencement  of 
the  operations,  the  small  air-cock  e  (Fig.  78)  was  used  to  permit 
the  continuous  admittance  of  air  to  the  suction  pipe  of  the  pump. 
This  practice,  however,  was  soon  abandoned,  because  the  agita- 


HYDROMETALLURGY   OF  SILVER 


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d,  d,  di  d.  d.  d,  f 


CYANIDATION  OF  AURIFEROUS  SILVER  ORES  317 

tion  was  seriously  affected  by  it.  The  entrance  of  air  into  the 
suction  pipe  had  a  detrimental  influence  on  the  capacity  of  the 
pump,  and  the  effect  was  found  to  be  injurious  to  the  best  agita- 
tion. Perhaps  the  chief  trouble  was  due  to  the  rapid  rising  to  the 
surface  of  the  imprisoned  air  immediately  on  being  expelled  from 
the  discharge-pipe.  The  air  bubbles  breaking,  on  reaching  the 
surface  of  the  charge,  caused  a  splendid  surface  movement  that 
might  be  easily  mistaken  for  the  thorough  agitation  of  the  entire 
mass  without  effecting  a  proper  scouring  of  the  bottom  point  of 
the  vat.  The  present  practice  is  to  allow  the  entrance  of  a  smaller 
quantity  of  air  into  the  mass,  through  the  small  air-cock  I  (Fig.  78). 

Ordinarily  a  charge  is  agitated  in  the  vats  from  forty  to  forty- 
four  hours,  after  which  it  is  discharged  into  the  corresponding 
decantation  vat,  where  it  is  usually  given  a  two  days'  treatment. 
Should  the  charge  from  the  agitation  vat  not  fill  the  decantation 
vat,  enough  precipitated  solution  is  pumped  up  from  the  strong 
solution  sump  to  fill  it;  after  agitation  for  half  an  hour,  the  charge 
is  allowed  to  settle.  Should  the  addition  of  this  extra  solution 
be  unnecessary,  the  charge  is  not  agitated,  but  allowed  to  settle 
as  long  as  practicable,  the  clear  supernatant  solution  being  mean- 
while decanted  off.  After  the  first  settling  and  decantation,  the 
vat  is  pumped  full  of  weak,  precipitated  solution,  which  is  usually 
of  a  strength  approximating  0.1  per  cent,  of  KCN  per  ton,  and 
the  charge  is  agitated  for  an  hour  or  two  by  means  of  the  3-in. 
centrifugal  pump  connected  with  the  vat,  about  25  Ib.  of  slaked 
lime  being  added  during  the  agitation.  The  pump  is  then  stopped 
and  an  additional  quantity  of  slaked  lime,  usually  about  10  Ib.,  is 
sprinkled  evenly  over  the  top  of  the  charge.  After  settling  a 
few  hours,  the  decantation  pipe  is  lowered  and  the  settling  and 
decanting  of  clear  solution  continued  as  long  as  practicable.  As 
many  washes  and  decantations  as  possible  within  the  time  limit 
of  the  treatment  are  given  in  this  manner.  When  permissible, 
the  last  wash  given  is  of  clear  water,  though  a  few  of  the  charges 
have  to  be  washed  entirely  with  weak  solution. 

When  treating  charges  containing  the  equivalent  of  15  tons  of 
dry  slime,  usually  four  settlings  and  four  decantations  can  be 
effected  with  the  forty-eight  hours  of  treatment,  each  decanta- 
tion averaging  about  22  tons  of  solution;  hence  about  90  tons  of 
solution  is  decanted  in  treating  a  15  ton  charge,  and  each  decan- 
tation removes  approximately  58  per  cent,  of  the  total  solution 


318 


HYDROMETALLURGY  OF  SILVER 


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sg 


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c3  r:    —  C-  r.  oj  CO         D.  d 

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fl.i 


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Qt  ^3  fs^    c3    O    rt 

$  J^  ^-S  E 


CYANIDATION  OF  AURIFEROUS  SILVER  ORES  319 

present.  Assuming  the  wash-agitation  to  be  perfect,  the  four 
decantations  should  then  theoretically  contain  about  97  per  cent, 
of  the  total  value  dissolved  at  the  time  the -washing  was  com- 
menced. The  settled  pulp  is  discharged  through  the  bottom 
valve  and  the  4-in.  discharge-pipe  into  the  waste  launder. 

Table  III  shows  the  rate  of  settling  per  hour,  determined  at 
various  times  on  several  different  charges. 

The  pulp,  ready  for  discharging,  carries  50  per  cent. of  moisture, 
the  contained  solution  averaging  0.07  per  cent,  of  KCN,  and  hav- 
ing an  average  value  of  approximately  $0.40  of  gold  and  1.50  oz. 
of  silver  per  ton.  These  values  are  higher  than  would  be  expected 
to  remain  in  the  solution  after  the  several  decantations  and  dilu- 
tions effected;  yet  (as  has  already  been  recorded  by  several  dif- 
ferent parties  operating  similar  slime  plants)  the  solution  of  value 
from  the  slime  does  not  cease  at  the  completion  of  the  agitation 
proper,  but  continues  throughout  the  washing;  the  value  of  the 
wash-solution  is  thus  being  constantly  augmented.  This  feature, 
however,  is  more  noticeable  with  the  silver  than  with  the  gold, 
and  the  maximum  extraction  of  the  gold  is  obtained  earlier. 
For  these  reasons,  the  solution  contained  in  the  discharged  pulp 
will  always  carry  more  values  than  it  should  according  to  calcula- 
tion based  solely  upon  successive  dilution  and  assuming  the  agi- 
tation to  be  perfect. 

A  portion  of  the  sample  of  the  pulp  ready  for  discharging 
(together  with  its  proper  proportion  of  contained  solution)  is 
dried,  the  assay  results  being  taken  to  represent  the  value  of  the 
discharged  slime.  Another  portion  of  the  pulp  is  washed  and 
then  assayed.  On  an  average  the  washed  sample  will  run  about 
$0.40  of  gold  and  from  1  to  2  oz.  of  silver  per  ton  lower  than  the 
unwashed  sample. 

The  3-in  centrifugal  pumps  connected  with  the  decantation  vats 
are  the  ordinary  pumps  commonly  used  for  pumping  solutions; 
the  only  alterations  being  that  the  bearing  nearest  the  pump 
shell  is  tapped  with  a  ^-in.  pipe,  which  supplies  the  bearing  with 
solution  under  pressure.  These  pumps  run  about  four  hours  in 
each  twenty-four,  and  have  given  excellent  satisfaction,  the  only 
repair  work  being  an  occasional  repacking  of  some  of  the  stuffing- 
boxes. 

Tables  IV  and  V,  giving  a  somewhat  detailed  record  of  the 
treatment  of  one  charge,  may  be  taken  to  represent  the  usual 


320 


HYDROMETALLURGY   OF  SILVER 


practice.  The  usual  charge  is  now  but  15  tons  of  slime  (dry 
weight),  while  the  proportion,  by  weight,  of  solution  to  slime  has 
been  increased  to  2.5  :  1. 

Precipitation.  —  All  solution  leaving  the  decantation  vats  is 
passed  through  the  zinc-boxes  before  being  reused.  The  zinc- 
boxes  have  to  be  watched  closely;  owing  to  the  excess  of  lime 
present  in  the  solution,  difficulty  is  experienced  in  obtaining  good 
precipitation.  Records  are  kept  of  the  quantity  of  solution  daily 
passing  through  the  boxes,  together  with  the  assay  values  of  the 
solution  before  and  after  precipitation.  These  records  show 
that,  during  the  last  three  months,  an  average  of  practically  48  tons 
of  strong  and  117  tons  of  weak  solution,  or  a  total  of  165  tons, 
was  passed  through  the  boxes  daily;  the  average  assays  of  the 
solution  were  as  given  herewith: 

TABLE  VI 


STRONG  SOLUTION 

WEAK  SOLUTION 

GOLD 

SILVER 

.     GOLD 

SILVER 

Entering  zinc-boxes 

$1.05 
0.10 

Oz. 

2.90 
0.40 

$0.60 
0.10 

Oz. 

1.70 
0.35 

Leaving  zinc-boxes  

The  zinc-boxes  have  a  combined  total  shavings  capacity  of 
approximately  120  cu.  ft. ;  the  rate  of  flow  of  the  solution  through 
the  boxes  during  1904  averaged  1.37  ton  per  cu.  ft.  of  shavings 
per  twenty-four  hours. 

The  highest-grade  precipitate  yet  recovered  from  the  slime 
plant  assayed  approximately  $6800  gold  and  17,300  oz.  of  silver 
per  ton. 

Tonnage,  Percentages,  etc.  —  The  normal  capacity  of  the  plant, 
while  treating  15-ton  charges,  and  allowing  a  two  days'  treatment 
in  both  agitation  and  decantation  vats,  is  30  tons  per  day.  Dur- 
ing the  last  quarter  of  1904  approximately  2550  tons  of  slime  (net 
dry  weight)  was  treated,  and  the  extraction  during  this  period 
(shown  by  the  differences  between  assays  of  the  charge  and  of  the 
residue)  was  74.9  per  cent,  of  the  gold  and  49.2  per  cent,  of  the 
silver.  During  this  period  the  assay  value  of  the  slime  averaged 
$4.35  of  gold  and  19.25  oz.  of  silver.  During  the  last  two  months 
(March  and  April,  1905)  3.56  Ib.  of  sodium  cyanide  (equivalent  to 


CYAN1DATION  OF  AURIFEROUS  SILVER  ORES  321 

4.40  Ib.  of  potassium  cyanide)  was  used  per  ton  of  slime  treated. 
The  average  extraction  of  silver  for  the  last  three  months  has 
been  51  per  cent.  The  consumption  of  cyanide,  zinc  and  slime 
per  ton  of  dried  slime  treated  during  this  time  was:  Sodium  cy- 
anide, 4.42  Ib.;  zinc,  0.957  Ib.,  and  lime,  13.95  Ib.  The  sodium- 
cyanide  consumption  is  equivalent  to  5.52  Ib.  of  potassium  cy- 
anide. 

Table  VII  gives  the  operating  costs  per  ton  of  slime  treated. 

TABLE  VII.  —  SLIME  COSTS  PER  TON 

Cyanide  (4.42  Ib.  @  $0.63)    .                           $2.785 

Zinc  (0.957  Ib.  @  $0.30) 0.287 

Lime  (13.95  Ib.  @  $0.0118)   0.165 

Other  supplies    0.238 

Lubricating 0.033 

Labor 0.491 

Salaries    0.748 

Assay  office  (labor  and  supplies) 0.066 

Power  (ditch  maintenance  and  supplies)    0.621 

Miscellaneous  (lighting,  etc.)   0.002 

Management  and  general  expenses 0.179 

Total  i $5.615 

CYANIDING  AURIFEROUS  SILVER  ORES  AT  SAN   SALVADOR,   C.    A. 

The  following  interesting  notes  on  the  working  of  this  process 
in  San  Salvador  were  communicated  to  the  Engineering  and 
Mining  Journal,  June  1,  1905,  by  Alfred  Chiddey: 

The  process  was  first  applied  to  tailings  from  the  pan  amal- 
gamation in  which  the  ore  was  first  roasted  with  salt  in  the  ordi- 
nary way.  After  all  these  tailings  had  been  worked,  the  process 
was  applied  to  raw  ore,  with  results  so  successful  that  roasting 
and  amalgamation  are  now  discontinued,  and  much  poorer  ore 
can  be  treated.  The  silver  occurs  presumably  as  sulphide  and 
the  ore  contains  a  little  copper,  which  apparently  helps  the  ex- 
traction; at  any  rate,  it  seems  to  do  no  harm. 

The  mode  of  working  is  as  follows:  The  ground  ore,  after 
leaving  the  arrastras  (which  are  fitted  with  30-mesh  screens), 
is  passed  over  amalgamated  plates  and  allowed  to  settle  in  masonry 

1  $5.615  Mexican  currency  during  this  period  was  equivalent  to  $2.66  gold. 
The  cost  of  realization  on  the  precipitate  is  not  included  in  the  working  cost: 
these  expenses  are  high.  The  average  cost  of  realization  on  the  precipitate 
produced  in  the  slime  plant,  per  ton  of  dry  slime,  is:  Government  taxes, 
$0.856;  treatment  charges  (including  transportation  expenses),  $1.202; 
total,  $2.058. 


322  HYDROMETALLURGY   OF  SILVER 

tanks,  where  a  rough  classification  is  effected  into  sand  and 
slimes.  The  sand  is  charged  over  gratings  with  25  Ib.  lime  per 
ton  into  the  percolation  vats.  A  cyanide  solution  of  0.40  per 
cent,  strength  is  introduced,  equivalent  to  nearly  one-half  the 
weight  of  the  sand.  After  standing  twelve  hours,  the  cock  is 
opened  and  the  vat  allowed  to  drain,  the  solution  passing  to  the 
zinc-boxes.  On  the  following  day  the  surface  is  leveled  off  and 
raked  over.  The  charge  is  then  allowed  to  stand  for  six  days 
exposed  to  the  air,  without  adding  more  solution.  On  the  expira- 
tion of  six  days,  a  0.20  per  cent,  solution  is  added,  and  the  leach- 
ing is  continued  rapidly  without  intermission  for  four  days,  at 
the  end  of  which  time  a  water-wash  is  employed. 

The  first  wash  that  comes  off,  after  the  vat  has  been  standing 
dry  for  six  days,  often  contains  from  12  to  20  oz.  silver  per  ton, 
which  will  run  down  the  second  day  to  6  oz.,  on  the  third  day  to 
3  oz.,  on  the  fourth  day  to  1^  oz.;  the  water-wash  will  generally 
be  from  0.5  to  1  oz.  per  ton.  The  sand  contains  from  13  to  15  oz. 
silver  and  from  $3  to  $5  gold  per  ton  before  treatment.  •  The 
extraction  for  the  past  18  months  has  averaged  from  85  to  90  per 
cent,  of  the  silver  and  90  to  92  per  cent,  of  the  gold. 

The  consumption  of  cyanide  is  a  little  under  two  Ib.  per  ton. 
Lately  sodium  cyanide  has  been  used  with  good  results.  It  is  a 
little  cheaper,  but  much  more  deliquescent  than  potassium  cy- 
anide, and  during  the  rainy  season  a  box  has  to  be  used  up  im- 
mediately it  is  opened. 

The  ore  contains  argentite,  chalcopyrite,  sometimes  specks 
of  fahlerz  (tetrahedrite),  but  the  amount  of  copper  is  small, 
generally  under  one-tenth  per  cent.  The  sump  solutions  always 
contain  copper,  but  presumably  as  sulphocyanate.  The  precipi- 
tation of  gold  and  silver  in  the  zinc-boxes  is  practically  perfect. 

The  slime  is  air-dried  and  mixed  with  the  sand  in  proportion 
of  1  :  2,  but  the  time  of  treatment  is  prolonged  to  18  days  in- 
stead of  12,  the  time  required  for  the  sand.  The  air-dried  slime 
is  previously  put  through  a  disintegrator.  This  method  of  treat- 
ing the  slime  is  only  temporary.  Although  the  extraction  is 
satisfactory  the  time  of  treatment  is  too  prolonged,  and  the 
method  is  available  only  during  the  dry  season. 

The  following  are  results  on  a  few  charges  of  sand: 


CYANIDATION  OF  AURIFEROUS  SILVER  ORES 


323 


No.  OF 
CHARGE 

BEFORE  LEACHING 

TAILINGS  AFTER  LEACHING 

SILVER  Oz. 

GOLD  Oz. 

SILVER  Oz. 

GOLD  Oz. 

44 

20.17 

0.37 

2.52 

0.06 

45 

15.50 

0.30 

0.95 

0.05 

46 

14.16 

0.30 

1.57 

0.03 

47 

16.50 

0.36 

1.87 

0.35(?) 

48 

13.10 

0.30 

1.44 

0.03 

49 

16.00 

0.30 

1.37 

0.03 

50 

14.70 

0.30 

1.37 

0.03 

55 

15.50 

0.26 

1.50 

0.02 

56 

13.70 

0.23 

1.00 

0.02 

57 

13.90 

0.21 

1.00 

0.02 

TREATMENT  OF  ROASTED  ORE 

As  stated  above,  complex  silver  ores  heavily  charged  with 
metal  sulphides  are  not  suitable  to  be  treated  raw  with  a  cy- 
anide solution  on  account  of  the  decomposing  action  of  these 
sulphides  on  potassium  cyanide.  To  make  this  process  possible, 
such  ores  have  first  to  be  roasted  with  salt  to  reduce  the  con- 
sumption of  cyanide.  The  roasting  has  to  be  done  as  carefully 
as  for  the  sodium  hyposulphite  process,  in  order  to  secure  a  mini- 
mum loss  of  silver  by  volatilization.  This  is  accomplished  by 
roasting  at  a  very  low  heat  without  even  raising  the  temperature 
toward  the  end,  so  that  as  few  of  the  metal  chlorides  are  expelled 
as  possible.  These  metal  chlorides,  if  expelled  by  heat,  are  the 
sole  cause  of  the  loss  of  silver  by  volatilization  (silver  chloride 
as  such  not  being  volatile),  and  it  would  be  folly  to  employ  this 
means  to  remove  them  from  the  ore.  It  is  much  more  rational  to 
remove  them  by  leaching  with  water,  though  the  resulting  base- 
metal  solution  will  have  to  undergo  a  treatment  similar  to  that 
practised  in  the  lixiviation  with  sodium  hyposulphite,  to  recover 
the  silver  dissolved  therein.  Having  the  ore  prepared  for  ex- 
traction as  far  as  that,  it  seems  to  be  more  rational  to  extract  the 
silver  chloride  and  the  gold  subchloride  first  with  sodium  hypo- 
sulphite, and  then,  after  the  latter  has  been  replaced  in  the  ore 
by  water,  to  apply  the  solution  of  potassium  or  sodium  cyanide 
to  extract  the  remainder  of  the  gold  and  that  part  of  the  silver 
which  was  not  chloridized.  This  procedure  is  advisable  and  much 
improved  extraction  can  be  thus  expected,  for  the  following 
reasons : 

It  will  make  the  application  of  the  cyanide  solution  possible 


324  HYDROMETALLURGY   OF  SILVER 

to  a  much  larger  variety  of  ores  than  if  the  same  is  applied  directly 
after  roasting  and  washing  of  the  ore.  Not  all  the  metal  salts 
formed  in  chloridizing  roasting  are  soluble  in  water;  some  will 
remain  in  the  ore.  like  cuprous  chloride,  lead  sulphate  and  others, 
and  will  act  decomposingly  on  the  cyanide  solution.  Most  of 
such  salts,  however,  dissolve  in  a  solution  of  sodium  hyposulphite, 
and  are  removed  from  the  ore  to  a  great  extent  during  silver 
leaching,  so  that  the  cyanide  solution  will  thus  be  applied  to  a 
much  cleaner  ore,  which  may  make  possible  the  treatment  of 
ores  which  contain  a  larger  percentage  of  copper,  lead  and  other 
metals.  The  action  of  the  cyanide  solution  will  be  more  energetic 
on  those  particles  which  escaped  chlorination,  and  suffer  less  by 
decomposition,  if  the  main  bulk  of  the  silver  and  other  metals 
has  been  removed.  This  is  also  the  case  with  regard  to  the  gold. 
It  will  be  found  that,  in  the  case  of  a  silver  ore  rich  enough  in 
gold  to  show  after  roasting  with  salt  a  perceptible  amount  pf 
gold  when  concentrated  in  a  horn  spoon,  the  gold  assumes  a 
much  brighter  color  if  this  is  done  with  a  hyposulphite  solution. 
In  treating  rich  silver-gold  concentrates,  as  related  in  Chapter 
XXII,  I  experienced  the  fact  that  by  first  roasting  with  salt,  then 
extracting  the  gold  by  Plattner's  method,  and  finally  the  silver 
with  sodium  hyposulphite,  only  50  per  cent,  of  the  gold  could  be 
extracted,  while  by  leaching  the  silver  first  and  then  applying 
the  Plattner  method,  the  gold  extraction  was  as  high  as  95  per 
cent.  There  is  no  reason  to  doubt  that,  from  an  auriferous  silver 
ore  which  was  first  chloridized,  the  cyanide  solution  will  extract 
the  gold  more  quickly  and  more  completely  if  the  ore  be  first 
leached  with  sodium  hyposulphite. 

The  increase  in  cost  will  be  but  very  slight,  as  it  will  not  in- 
volve any  handling  of  the  charge,  but  will  require  only  a  continua- 
tion of  the  leaching  with  another  solvent.  The  total  time  may 
even  prove  to  be  shorter  than  if  only  a  cyanide  solution  was  used. 

It  seems  that  very  high  extractions  of  silver  and  gold  may 
be  obtained  by  a  combination  of  the  two  processes,  and  thorough 
experimenting  on  that  line  will  in  all  probability  be  well  rewarded. 
This  refers,  of  course,  only  to  higher  grade  ores  that  can  stand 
the  cost  of  roasting,  and  the  nature  of  which  resists  a  high  chlori- 
nation. 

John  F.  Allan,  City  of  Mexico,  in  a  very  interesting  paper  on 
"Cyanide  Treatment  of  Silver  Ores  in  Mexico,"  read  before  the 


CYANIDATION  OF  AURIFEROUS  SILVER  ORES  325 

Atlantic  City  meeting,  February,  1904,  of  the  American  Institute 
of  Mining  Engineers,  gives  two  examples  of  cyaniding  silver  ores 
containing  gold  which  were  roasted  with  salt  before  leaching: 

"Leaching:  Example  No.  1. — The  plant  in  question  has  a 
capacity  of  700  tons  per  month.  The  chloridized  pulp  is  charged 
into  filters  18  ft.  in  diameter  and  4  ft.  deep,  and  receives  as  pre- 
liminary treatment  four  hours'  soaking  with  water  and  seven  hours' 
percolation.  This  water-wash  is  passed  through  special  zinc- 
boxes,  where  an  impure  precipitate  containing  gold  and  silver 
is  formed,  and  a  consumption  of  about  0.4  Ib.  of  zinc  per  ton 
takes  place.  When  the  salt  and  impurities  have  been  removed 
by  the  wash-water,  the  charge  receives  fifteen  hours'  soaking  and 
ninety  hours'  percolation  with  cyanide  solution  of  0.3  percent, 
or  6  Ib.  to  the  ton  of  water,  the  ore  receiving  an  equivalent  of  1.5 
of  solution  to  1  of  ore,  or  0.45  per  cent.  To  displace  the  strong 
solution,  twelve  hours  of  percolation  with  a  0.05  per  cent,  weak 
solution  is  given,  in  the  proportion  of  1  of  water  to  2  of  ore,  and 
a  final  thirty  hours'  percolation  with  water-wash.  The  con- 
sumption of  cyanide  is  2.15  Ib.  per  ton,  and  zinc  1.08  Ib. 

"The  extractions  are:  Gold,  76;  silver,  85.25;  total  value,  81.77 
per  cent.  It  will  be  observed  that  the  gold  extraction  is  not  as 
good  as  the  silver,  a  fact  often  noted  in  chloridized  ores.  The 
following  are  the  costs,  which  can  be  considered  as  typical  in  small 
plants,  although  in  some  instances  they  have  been  reduced: 
Superintendence,  $0.62;  labor,  $0.96;  cyanide,  $1.71;  zinc,  $0.29; 
laboratory,  $0.10;  various  stores,  $0.50;  total,  $4.18  Mex.,  or, 
at  50c.  U.  S.  per  $1  Mex.,  $2.09  U.  S.  per  ton." 

"Leaching:  Example  No.  2.  —  The  plant  is  capable  of  treat- 
ing 1800  tons  per  month.  The  treatment,  with  small  variations, 
is  practically  the  same  as  in  Example  No.  1.  The  consumption 
of  cyanide  is  1.73  Ib.  and  of  zinc  1.27  Ib.  per  ton,  the  high  consump- 
tion of  zinc  being  due  to  the  ore  averaging  over  30  oz.  of  silver 
per  ton.  The  extractions  are:  Gold,  82.30;  silver,  77.32;  total 
value,  79  per  cent.  The  costs  are:  Superintendence,  $0.33; 
labor,  $0.31;  cyanide,  $0.99;  zinc,  $0.33;  laboratory,  $0.10;  vari- 
ous stores,  $0.22;  total,  $2.28  Mex.,  or,  at  50c.  U.  S.  per  $1  Mex., 
$1.14  U.  S.  per  ton." 

It  is  to  be  regretted  that  no  analysis  or  description  of  the  ore 
is  given.  The  fact  that  in  Example  No.  1  the  gold  extraction 
is  only  76,  and  the  silver  85.25  per  cent.,  and  Mr.  Allan's  remark 


326  HYDROMETALLURGY  OF  SILVER 

that  the  fact  of  a  lower  gold  extraction  is  often  noted  in  chlori- 
dized  ores,  verify  my  experience  in  treating  auriferous  silver 
ores.  If  to  the  roasted  ore  of  Example  No.  1  a  solution  of  sodium 
hyposulphite  had  been  applied  before  the  cyanide  solution,  the 
extraction  of  both  metals  would  have  been  higher,  and  very 
likely  that  of  gold  better  than  the  extraction  of  the  silver. 

TESTING  THE  CYANIDE  SOLUTION  FOR  GOLD  AND  SILVER 

In  order  to  conduct  the  cyanide  process  intelligently  frequent 
tests  of  the  solutions  for  gold  and  silver  are  necessary. 

Arents'  Test. l  —  This  test  is  based  upon  the  fact  that  metallic 
copper  will  precipitate  gold  and  silver  upon  its  surface  from 
acid  solution.  Of  course  the  fact  is  not  new,  but  its  application 
is  probably  so.  Arents  has  used  the  method  with  success;  it 
recommends  itself  by  the  rapidity  and  ease  with  which  it  may  be 
carried  on. 

An  auriferous  cyanide  solution,  if  made  acid  with  sulphuric 
acid  and  boiled  with  finely  divided,  pulverulent,  metallic  copper, 
will,  within  a  short  time,  deposit  its  gold  content  on  the  copper. 
Any  silver  in  the  solution  is  also  precipitated.  If  this  mixture 
is  now  filtered,  the  filter  and  contents  may  at  once  be  subjected 
to  a  crucible  assay  treatment,  and  its  lead  button  cupeled  and 
determinated. 

If,  instead  of  taking  cement  copper,  or  any  metallic  copper 
powder,  a  solution  of  bluestone  is  used  after  acidification,  and  a 
few  small  pieces  of  sheet  aluminum  are  added,  and  the  solution 
boiled  until  all  the  copper  has  come  down,  the  result  as  to  the 
precipitation  of  gold  and  silver  is  the  same.  This  modification 
takes  more  time  and  attention  in  boiling.  If  aluminum  has 
been  used,  it  should  go  into  the  crucible  with  the  filter  and  its 
contents.  Commercial  cement  copper  is  particularly  fitted  for 
this  test,  because  the  acid,  in  taking  up  any  basic  iron  or  copper 
salts  of  the  cement  copper,  renders  the  copper  as  finely  divided 
as  it  is  customary  to  obtain  in  the  sluice-boxes  of  copper  leachers. 
The  finer  and  the  more  pulverulent  the  copper  is,  the  greater  is  its 
surface  and  the  more  energetic  the  precipitation,  thus  permitting 
a  minimum  amount  of  copper  to  be  used. 

In  applying  the  method,  Arents  uses,  as  a  rule,  250  c.c.  of  the 

1  From  a  paper  by  Albert  Arents  read  before  the  Albany  meeting,  Feb. 
1903,  of  the  American  Institute  of  Mining  Engineers. 


CYANIDATION  OF  AURIFEROUS  SILVER  ORES  327 

solution  to  be  tested  and  a  few  c.c.  of  sulphuric  acid,  agitates  for 
several  seconds,  and  then  adds  not  less  (although  not  much  more) 
than  one  gram  of  cement  copper.  Now  follows  heating  to  boil- 
ing. This  is  kept  up  for  about  10  minutes,  so  that  the  rising 
steam-bubbles  keep  the  mixture  well  agitated.  The  mixture  is 
then  filtered  through  a  7-in.  diameter  gray  filter-paper.  No 
washing  is  done.  As  soon  as  the  filtering  is  finished,  one-third  of 
a  crucible  charge  of  flux  is  added  to  the  filter  containing  all  the 
sediment  of  the  mixture.  Some  of  the  moisture  is  rapidly  absorbed 
by  the  flux,  which  permits  the  folding  of  the  filter's  rim  upon  the 
charge  and  its  subsequent  removal  without  loss  or  tearing.  One- 
third  of  a  crucible  charge  of  flux  having  previously  been  placed 
upon  the  bottom  of  the  crucible  which  is  to  be  used  for  melting, 
the  filter  is  transferred  to  the  crucible,  well  tucked  down,  and  the 
last  third  of  the  crucible  charge  is  placed  on  top  of  the  filter  in  the 
crucible.  It  is  then  ready  for  the  furnace.  The  filter  itself  fur- 
nishes the  reducing  agent  for  the  assay.  Arents  uses  30  grams 
litharge  and  the  usual  amount  of  borax  and  soda,  employing  a 
No.  F  crucible  for  melting.  About  20  grams  of  lead  are  obtained. 
The  lead  button  comes  out  bright  and  clean,  and  upon  cupeling 
furnishes  a  bead  free  from  copper. 

Possibly  this  method  of  testing  for  gold  and  silver  may  be 
used  upon  other  solutions  than  cyanide;  also,  for  solutions  from 
testing  metallic  copper  for  precious  metals,  when  the  solutions  do 
not  contain  nitric  acid  in  any  form. 

Alfred  Chiddey's  Test.  —  Four  assay  tons  of  solution  are  taken 
in  a  porcelain  dish,  and  20  c.c.  of  a  10  per  cent,  solution  of  lead 
acetate  added;  then  4  grams  of  zinc  shaving  and  afterward  20  c.c. 
of  hydrochloric  acid.  When  the  action  has  nearly  ceased,  the 
contents  of  the  dish  are  boiled  for  a  minute  or  two  and  filtered. 
The  precipitate  is  well  washed  with  water,  moistened  with  alco- 
hol, dried,  wrapped  with  the  filter  in  lead  foil,  and  cupeled. 

The  testing  of  the  strength  of  a  solution  in  potassium  cyanide 
is  usually  done  by  titrating  the  same  with  a  standard  solution 
of  silver  nitrate.  This  standard  solution  is  prepared  by  dissolv- 
ing 13.04  grams  of  c.p.  silver  nitrate  in  one  liter  of  distilled  water. 
Every  0.1  c.c.  of  such  a  solution  added  to  10  c.c.  cyanide  solu- 
tion is  equivalent  to  0.01  per  cent,  of  potassium  cyanide.  The 
operations  are  as  follows: 

A  Mohr's  burette,  graduated  to  0.1   c.c.,  is  filled  with  the 


328  HYDROMETALLURGY  OF  SILVER 

standard  silver  solution.  Ten  c.c.  of  the  cyanide  solution  which 
is  to  be  tested  is  taken  up  with  a  pipette  and  emptied  into  a  small 
flask  which  is  brought  under  the  burette.  Then  drop  by  drop  the 
silver  solution  is  added.  Each  drop  produces  a  whitish  precipi- 
tate, which,  however,  disappears  again  when  the  flask  is  well 
shaken.  This  is  continued  until  the  last  drop  produces  a  cloudi- 
ness or  turpidity,  which  remains  even  after  a  vigorous  shaking  of 
the  flask. 

The  scale  on  the  burette  is  now  read.  Each  tenth  of  a  c.c. 
silver  solution  used  indicates  0.01  per  cent,  of  potassium  cyanide 
contained  in  the  solution. 


INDEX 


Agitating  tanks 263 

Agitation  vats 304,  305 

Agitators,  in  precipitation  vats.    185 

Air  blow-off  drum 200 

compressed,   as   agitator   in 

precipitation  vats 185 

effect  of,  in  roasting  calca- 
reous ore 147 

with    highly    sulphureted 

ore    134 

provision   for,    in    Bruckner 

furnace 69 

required       in       chloridizing 

roasting 4,  56 

Allan,  John  F '.'.  .324,  325 

Alumina,  causes  loss  of  silver  in 

roasting 8 

Aluminum   chloride,    formed   in 

roasting 8 

Amalgamation Pref .   iii 

barrel 37 

pan 37 

roasting  for  ....     37 
volatile  chlorides 
undesirable  in      20 

American  underfed  stoker 272 

Analysis  of  ores  of  Anglo-Mex- 
ican Mining  Co., 

Yedras,  Mexico 127 

of  precipitate  of  ores 

of  A  vino,  Mexico  ...    195 
of   San   Francisco   del 

Oroore    100,118 

Anglo-Mexican  Mining  Co.,  Ye- 
dras, Sinaloa,  Mexico,  analysis 

of  ores  of    127 

roasting  ores  of  ....    50 
Anhydrous  sulphuric  acid,   for- 
mation of 4 


Antimonial  fahlerz,  effect  of 
steam  on  ores  con- 
taining    33 

galena,   effect   of 
steam  on  ores  con- 
taining       33 

minerals,  effect  of,  on 

time  of  lixiviation   180 
silver  minerals,    be- 
havior of,  in  roast- 
ing           9 

Antimony  antimonate,  formed  in 

roasting 7 

expelled  by  steam  in 

roasting 31 

sulphide,  behavior  of, 

in  roasting 7 

trichloride,  action  of, 

in  roasting 7 

Appearance  of  ore  in  steps  of 

roasting 5 

Arch,    construction    of    in    long 

reverberatory  furnace 55 

Arents,  Albert 326,  327 

Arents'  test  for  gold  and  silver.   326 
Argentiferous  black  copper,  ex- 
traction of  silver 

from 278 

copper  matte,  used 
for     sulphating 

roasting 94 

zinc-lead  ore,  chlo- 
ridizing of   ....     99 
Argentite,  behavior  of ,  in  roasting       9 
Arsenate    of   silver,    formed    in 

roasting. . .       7 
soluble  in  so- 
dium hypo- 
sulphite  . .       7 


329 


330 


INDEX 


Arsenic,  action  of,  in  roasting. .       7 
chloride,      formed      in 

roasting 7 

expelled    by    steam    in 

roasting 31 

Arsenical  minerals,  effect  of  on 

time  of  lixiviation  . .   180 
ores,  at  Yedras,  Mex- 
ico, experi- 
ments with    . .   127 
long  furnace  for.     50 
pyrites,  behavior  of,  in 

roasting 7 

silver  minerals,  behav- 
ior of,  in  roasting. .       9 
Arsenious  oxide,  formed  in  roast- 
ing         7 

Assay  values  of  raw  and  roasted 
ores  of  Cusihuiriachic   Silver 

Mining  Company 24 

Augustin  process 155,  256 

Auric    chloride,    action    of,    in 

roasting 35 

Auriferous   black  copper,  extrac- 
tion   of    silver 

from 278 

silver  ores,  at  Pal- 

marejo,  Mexico  288 
Kiss  process  for. .  254 

roasting  of  34 

treatment,  Pref .  iv,  256 
Aurous  chloride,    action  of,    in 

roasting    35 

Avino,  Durango,  Mexico,  analy- 
sis of  precipitate  of  ores  of    . .   195 

Bag  system  of  collecting  dust  . .     91 

Ball-mill 143 

Balling  of  the  ore  142 

Barrel  amalgamation 37 

Base-metal  chlorides,  action  of, 
in     lixiviation 

troughs    222 

desilverizing     by 

water 170 

expulsion    of,    in 
amalgamation  38, 39 


Base-metal    solubility   of  silver 

chloride  in    ...  161 
see      also     Metal 
chlorides 

leaching 157 

at     Sombrerete, 

Mexico 166 

cause  of  chlorina- 

tion 115 

cupric      chloride 
.     used  during.  .167, 
173 

description  of  . .   156 
of  silver  ores  rich 

in  gold 284 

tanks  for 229 

time  required  for  244 
solutions,     loss     of 

silver  in  . .  162,  246 
precipitation      of 

silver  from  ...   164 
Battery,  stamp,  adding  salt  in.  .    140 
effect  of  crush- 
ing ore  in  ...      11 
Bins,  construction  of,  for  heap- 
roasting    31 

Bisulphide  of  calcium,  formed  in 

calcium  sulphide  tank    190 

Bituminous  coal,  used  in  chlori- 

dizing  roasting 42 

used  in  sulphating  roasting  . .     95 
Black  copper,  extraction  of  silver 

from 258,278 

Blake  rock-crusher 289 

Blue  vitriol,  formed   in  extrac- 
tion with  sulphuric 

acid 259,275 

in  base-metal  leach- 
ing at  Sombrerete  169 
Bosque  mill  at  Parral,  Mexico  101, 243 
Bottom    of    long    reverberatory 

furnace,  construction  of 52 

Brown  furnace 71 

Bruckner    63 

Bruckner      cylinders,  tests   for 

loss  of  weight  in .     23 
furnace    144, 147 


INDEX 


331 


Bruckner  furnace,  action  of  salt  in    18 
advantages  of    .  .     66 
advantages     over 
reverberatory 

furnace    30 

behavior  of  ore  in  140 
decomposition   of 

soluble  silver  in  146 
dust  formed  in  .  .  87 
effect  of  adding 

salt  in 135 

extra  handling  in  143 
Hofmann      im- 
proved       67 

loss  of  silver  in, 
less  than  in  re- 
verberatory 

furnace 20 

not  suited  for 
roasting  zinc 

lead  ores 126 

ores  suitable  for. .     67 
roasting    calcare- 
ous ore  in  ...    128 
roasting    sulphu- 

reted  ores  in  . .     42 
self-roasting  pro- 
cess in 152 

sulphureted    ores 
favorable       to 
roasting  in  ...    150 
used  at    Yedras, 

Mexico   17 

used  in  chloridiz- 

ing  self-roasting    26 
used     in    experi- 
ments with  ore 
from  Silver  King 

mine    34 

revolving  furnace    .     63 
speed  of  revolu- 
tion       66 

roaster 63 

Burlap,  used  for  filter  in  leaching 
tanks 183 

Cadmium,  in  zinc  blende  ore ...     99 


Calcareous  arsenical  silver  ore  at 
Yedras,  Mexico,  be- 
havior   of,    in 

roasting 17 

experiments  with  127 
loss  of    silver  in 

roasting 21 

ores,  chloridizing  . .  .    127 
long    furnace    for 

roasting 50 

Calcined  copperas,  added  in  oxi- 
dizing roasting 18 

Calcium  chloride  formed  in  roast- 
ing         8 

hyposulphite,  action  of, 

in  silver  leaching    .    179 
used  in  Kiss  process  .  254 
polysulphide,     prepara- 
tion of   186 

sulphate,  cause  of  ball- 
ing        143 

.     sulphide,  as  precipitant  164, 
179, 183, 198 
as  test  for  sil- 
ver     178 

boiler  and  pres- 
sure tank  for   188 
Calcspar,  in  gangue  of  ore  from 

San  Francisco  del  Oro  mine.   100 
Capacity  of  long  reverberatory 

furnace 60 

Carbonate  of  lime,  as  material 

for  cupel   .  .   210 
behavior  of,  in 

roasting  ....    8 
effect    of,    in 

roasting  ore  131 
used  to  decom- 
pose     base 
chlorides  . .     39 

Caustic  lime,  action  of,  in  roast- 
ing calcareous  ores.  .8,  134 
potash,  manufacture    of  205 
soda,  effect  of,  on  silver 

chloride 204 

Cement   copper,   as   precipitant 
for  silver 165, 167, 257 


332 


INDEX 


PAGE 

Centrifugal  pumps    304,  319 

Challenge  feeders 289 

used  in  the  Ropp 

furnace    77 

Charge  hopper,  construction  of, 

in  long  reverberatory  furnace     55 
Charging  a   long  reverberatory 

furnace 57 

Chemical  loss  in  weight    during 

roasting 20 

Chiddey,  Alfred 321,  327 

Chiddey's  test  for  gold  and  silver  327 
Chloride  of  copper,  used  to  cor- 
rect bad  roasting 104 

Chloridizing  heap-roasting 27 

advantages  of  . .  31 
of  calcareous  ores.  127 
period  of  roasting.  4 
roasting,  definition 

of Pref .  iii 

in      relation      to 
amalgamation 

Pref.  iii 

modification     of, 
for         lixivia- 

tion Pref.  iv 

of     argentiferous 
zinc-lead     ore, 
conclusions    of 
experiments    .   Ill 
experiments  ....   101 

theory  of 3 

with  steam    ....     31 
see  also  Roasting 

self-roasting 26 

in  Bruckner  revolving 

furnace 65 

Chlorination,  after  ore  has  left 

furnace 115 

aided     by     base- 
metal  leaching     115 
effect  of  salt  on  . .      16 

gold 36 

lowered  by  excess 

of  salt 107 

of  silver,  methods 
of  effecting   3 


Chlorination  on  cooling  floor    . .     27 

Chlorine,    formed    in     roasting     3, 

6, 8,  29,  30 

Clay,  es  material  for  cupel 210 

behavior  of,  in  roasting  . .       8 

Clean-up  box 301 

of  zinc-boxes  in  cyani- 

dation 300 

Cleaning  a  Bruckner  furnace. . .     71 
Coarse  crushing,  advantages  and 

disadvantages  of 12 

Compressed  air,  as  agitator  in 
precipitation 

vats 185 

in      pressure 

tanks 199 

Consolidated  Kansas  City  Smelt- 
ing  and    Refining   Company, 

Argentine,  Kansas    215, 261 

Continually  discharging  furnaces, 

fuel  needed  in    42 

Continuous   feeding  mechanical 

roasting  furnaces 71 

Cooling  floor,  Chlorination  on  ...     27 
Copper,  as  precipitant  for  silver  165 

effect  of,  in  ores 15 

in  silver  leach- 
ing        178 

on      time      of 

lixiviation    .    180 
matte,  extraction  with 
sulphuric  acid 

from 260 

in  Augustin  pro- 
cess    256 

suited  for  Zier- 

vogel  process  281 
pyrites,  behavior  of,  in 

roasting 6,  9 

in  ore  from  San 
Francisco  del 
Oro  mine  ...    100 
removing  from  precipi- 
tate before  cupellation  195 
sulphides,  advantage  of, 
inchloridiz- 
ing  ores   ...       4 


INDEX 


333 


Copper  sulphides,  necessary     in 
ores  for  sul- 
phating 
roasting. .  .     94 

Cost  of  cyaniding  auriferous  sil- 
ver ore  303 

of  roasting  in  the  modified 

Howell  furnace 119 

Crushing  in  stamp  batteries   ...     11 
ore  from  San  Francisco 
del  Oro  mine,  exper- 

ments  in 11 

through  rolls 11 

Cupeling  furnace,  dust-collector 

for 213 

refining     precipitate 

in 179,209 

Cupellation,  refining  precipitate 

by    195 

Cupric  chloride,  action     of,     in 

chloridizing  silver  . .       4 
as  precipitant  for  silver  168 
formed  in  roasting ....       6 
in    base-metal    leach- 
ing  167,173 

in    water    an    aid    to 

chlorination 115 

preparation  of 173 

treating  ore  with  ....    125 
used      in      base-metal 

leaching 173 

used  in  foul  hypo  solu- 
tion     181 

oxide,  action  of,  in  roast- 
ing           6 

sulphate,  formed  in  roast- 
ing           6 

formed    in    sulphating 

roasting 94 

product   of  extraction 

with  sulphuric  acid .   258 
Cuprous  chloride,  effect    of,    in 

leaching  . .   177 
formed         in 

roasting.  . .       6 
in  base-metal 
leaching  . .    168 


Cuprous    oxide,    action    of,    in 

roasting 6 

Cusihuiriachic,  Chihuahua,  Mex- 
ico    254 

lixiviation  at 155,  240 

loss  of   silver  in  leaching 

ore  from 163 

time  required  for  lixivia- 
tion at 180, 182 

working  ore  from 14 

Cusihuiriachic   Silver   Mining 

Company,  Chihuahua,  Mexico  24, 

80,  221 

Cyanidation  of  sand 292 

Cyanide  leaching  plant 293 

leaching  vats 292 

process  for  silver  ores 

Pref.  iv. 
Cyaniding  auriferous  silver  ores, 

at  Palmarejo,  Mexico . .  288 

at  San  Salvador,  C.  A.  .  321 
consumption  of  cyanide, 

zinc,  and  lime 302 

description      of      slime 

plant 305 

in  Mexico 325 

precipitation 320 

precipitation    of    silver 

and  gold 299 

sizing  test  on  slime.  .  .  .  305 

time  required    317 

tonnage  and  extraction .  302 
tonnage,      percentages, 

etc 320 

treatment  of  roasted  ore  323 

treatment  of  slime.  . . .  304 

working  costs   303 

Dead  roast 6, 18,  259 

Decantation  vats 304,  311 

Decrepitation,  caused  by  salt  in 

a  Bruckner  furnace 18 

Del  Oro  ore,  see  San  Francisco 

del  Oro  ore 

Desilverization  of  waste  liquor.  165 
Desilverizing    base-metal    chlo- 
rides with  water   .                   .170 


334 


INDEX 


Distillation  of  sulphur  from  pre- 
cipitate    207 

Distributing  trough  for  milk  of 

lime 188 

Don  Enrique  Mining  Company, 
Cusihuiriachic,    Chihuahua, 

Mexico 240 

Drum,  air  blow-off   200 

Drying  and  roasting  furnace  for 

silver  precipitate 208 

Dust  in  White-Howell    furnace     80 
collecting    methods,     bag 

system 91 

Hofmann's  flue-dust  col- 
lector    88,214 

Dusting  of  ore 27 

English  cupeling  furnace 209 

Experiments    in     roasting    ore 
from  San  Francisco  del  Oro 

mine    101 

Extraction,  of  the  gold  from  rich 

silver-gold  ores  . .  284 
with  sulphuric  acid.  258 

Fahlerz,  behavior  of,  in  roasting      6 
Ferric  arsenate,  formed  in  roast- 
ing           7 

chloride,    action    of,    in 

roasting 5,6,7 

oxide,  formed  in  roasting  6,  7 
sulphate,  formed  in  roast- 
ing      6,  7 

Ferrous  chloride,   action   of,    in 

roasting 5,6,7 

sulphate,  added  in  oxi- 
dizing roast- 
ing    18 

formed  in  sul- 
phating  roast- 
ing         94 

Filter  bottom  of  cyanide  leach- 
ing vats 292 

construction  of 159 

press 267 

Johnson 199 

Filtering  precipitate 204 


Filtering  quality  of  ore  improved 

by  lead  sulphide 14 

Filters  for  leaching  tanks 182 

Filtration,  effect   of   coarse  and 

fine  crushing  on    .  .      12 
effect  of,  on  time  of 

lixiviation    181 

Fine  crushing,  effect  of,  on  fil- 
tration    12 

Flint,  Idaho,  treatment  of  ores 

from « 40 

Flue-dust  collecting  methods  .  .     87 

collector,    Hofmann's   88, 

214 

-hole,  construction  of,  in 
long  reverberatory  fur- 
nace    55 

Free  filtration,  effect  of,  on  time 

of  lixiviation 180 

percolation,  conditions 

which  aid,  in  roasting.      12 
effect  of  coarse  and  fine 

crushing  on 12 

Freiberg,     Saxony,     extraction 

with  sulphuric  acid  at    258 

Fuel,  consumption  of,  in  roast- 
ing   32,42,119 

Furnace,  see  Bruckner,  Howell, 
Howell-White,  Long  reverber- 
atory, McDougal,  Mechanical 
roasting,  O'Harra,  Pearce,  Re- 
verberatory, Ropp,  Stetefeldt, 
Two-story  reverberatory 

Galena,  behavior  of,  in  roasting    6,  9 
effect    of,    on    time    of 

lixiviation   180 

effect    of    salt    on,    in 

roasting 16 

steam  in  roasting  ores 

containing 32 

in  ore  from  San  Fran- 
cisco del  Oro  mine  .  .     99 
not  desirable  in  a  charge       6 
not  favorable  in  roast- 
ing in  Stetefeldt  fur- 
nace..  86 


INDEX 


335 


Gangue,  effect  of,  on  free  perco- 
lation       12 

minerals        containing 
alumina,  behavior  of 

in  roasting 8 

of  ore  from  San  Fran- 
cisco del  Oro  mine  .    100 
Gas,  producer,  as  fuel  for  roast- 
ing       42 

Gay-Lussac  tower 91 

German  cupeling  furnace 209 

Gerstenhofer  pyrites  roaster  .  .  83 
Glauberite,  cause  of  balling  ....  143 
Globules  of  ore,  effect  of  steam 

on,  in  roasting 32 

Gold    chlorides,    action    of,    in 

roasting 35 

chlorination 36 

cyanidation  of  silver  ores 

rich  in    287 

extraction    of,  from    rich 

silver-gold  ores 284 

precipitation    of,   in   cya- 

niding 299 

roasting  of  silver  ores  con- 
taining       35 

testing    cyanide    solution 

for 326 

treatment    of   silver   ores 

rich  in    283 

Gravel,  as  filter  in  silver  leach- 
ing      182 

Gray  copper  ore,  behavior  of,  in 

roasting 6,  9 

Grinding  machine 238 

Grizzlies   289 

Gypsum,    removal   of,    in   Zier- 
vogel  process 282 

Hand-worked  furnaces 45 

reverberatory  fur- 
naces vs.  me- 
chanical roast- 
ing    62 

Hauch 155 

Heap-roasting   27 

advantages  of  . .     31 


Heat,  amount  of,  required  in  long 

reverberatory  furnace.  .     59 
excess  of,  cause  of  loss  of 

silver    20 

Hidalgo  Mining  Company,  Parral, 

Mexico    170, 171 

Hofmann    improved     Bruckner 

furnace 67 

Hofmann 's  flue-dust  collector  88,  214 
methods  of  extract- 
ing silver  with  sul- 
phuric acid    259 

modified  Howell  fur- 
nace        81 

Homestake  mortar 290 

Howell  furnace,  fuel  needed  in    42 
Hofmann 's  modified  .     81 

modified 243 

cost  of  roasting  in . .   119 
roasting  San  Fran- 
cisco del  Oro  ore 

in    113 

not  successful  in  roast- 
ing San    Francisco 

del  Oro  ore 112 

preferable  to  long  re- 
verberatory       52 

provision  for  air  in  .  .    128 
roasting  zinc-lead  ores 

in 112 

steam  used  in 110 

tests  made  in,  at  Chi- 
huahua, Mexico ...     24 
used  after  roasting  in 

Stetefeldt  furnace.    109 
used    at    Chihuahua, 

Mexico .  221 

Howell-White  furnace,  descrip- 
tion of   77 

dust-collector  in  ...88,91 

dust  formed  in 87 

good  fuel-economizer    44 
in     Bosque    mill    at 

Parral 101 

remedy  for  dust  in  . .     80 
see  also  White-Howell 
furnace 


336 


INDEX 


Hydrochloric    acid    formed    by 
steam  in 
Howell 
furnace .    1 10 
in  roasting    31 
formed  in  roast- 
ing   3,4,6 

used  in  test  for 
silver    183 

Iodine  solution,  test  for  silver. .   178 
Iron  chlorides  the  principal  chlo- 

ridizers 5 

effect  of,  in  ores 15 

pyrites,   advantage    of,    in 

chloridizing  ores        4 
behavior      of,      in 

roasting 5,9 

in    ore    from    San 
Francisco         del 

Oro  mine 100 

sulphate,  action  of,  in  roast- 
ing         17 

sulphide,  action  of,  in  roast- 
ing           6 

advantageous  in  sulphat- 

ing  roasting 94 

contained  in  zinc  blende       7 
sulphides  needed  in  chlori- 
dizing silver 5 

needed     in      sulphating 
roasting  ores 94 

Joachimsthal,  Bohemia,  lixivia- 

tion  at 155 

Johnson  filter  press 199 

Kiss 179, 254 

Kiss  process 254 

Krupp  ball-mill 261, 263 

Kustel,  G 39, 40, 41, 71, 86 

La  Baranca,  Sonora,  Mexico, 
gravel  used  for  filter  at  182 

La  Dura,  Sonora,  Mexico,  lixivi- 
ation  at 155 

Labor  required  on  long  reverber- 
atory  furnace 59 


Las  Bronzas,  Mexico,  lixiviation 

at 155 

Leach-troughs 215 

Leaching  auriferous  silver  ores  .   325 

base-metal 157 

at  Sombrerete 166 

description  of 156 

silver    174 

description  of   156 

method  of   177 

tanks,  construction  of     157 

filters  for 182 

to  remove   base-metal 

chlorides    21,40 

Lead  bath,  refining  precipitate  on  209 
carbonate,   by-product  in 

silver  leaching 178 

chloride,  action  of  in  amal- 
gamation ....     41 
formed  in  roast- 
ing           6 

effect  of,  in  ores 15 

in  silver  leaching 178 

lining  not  suited  for  lixivi- 
ation troughs    216 

oxide,  formed  in  roasting.       6 
silicate  formed  in  roasting    32 
sulphate,  action  of  in  amal- 
gamation       41 

effect  of,  in  leaching.  .    177 
formed  in  roasting  ....       6 
sulphide,  an  effective  agent 

in  filtering  ore 14 

see  also  Galena 
Lexington     mine,     experiments 

with  ore  from 28 

Lime  beneficial  to  roasting   ...       8 
milk  of,  as  precipitant  for 

silver    164 

used  in  preparing  cal- 
cium polysulphide . .  .    186 
rock,  see  Carbonate  of  lime 
Limestone,  cause  of  balling  ....   143 
gangue,  effect  of,  in 

roasting 10 

Litharge,  used  in  refining  silver 
precipitate 208 


INDEX 


337 


Lixiviation,  Augustin  process  .  256 
effect  of  antimonial 
minerals        on 

time  of 180 

arsenical  minerals 
on  time  of.  ...  180 

Kiss  process    254 

process,  reaction  of 
calcium  sul- 
phide and 
sodium  sul- 
phate   198 

tests  for  silver  in  178 
treatment  of  pre- 
cipitate in ....    198 
used  on  calcare- 
ous ores 127 

volatile  chlorides 
not  objection- 
able in 21 

Russell  process  ...   251 

tank 219 

time  required  for.  .    180 

trough    162,  219 

troughs 224 

arrangement  and 

operations  of. .  229 
construction  of.  .   215 
with  sodium  hypo- 
sulphite       155 

description        of 

process 156 

first     introduced 

Pref.  iv 

Long,  J.  T 171 

Long  reverberatory  furnace  ...     47 

capacity  of    60 

charging    57 

construction  of 52 

heat  required 59 

labor  required 59 

not  useful  for  roasting 
ores  low  in  sulphur      52 

two-story 60 

used  on  calcareous  ores 

at  Yedras,  Mexico..    128 
Loss  of  silver  by  volatilization  .   117 


Loss  of  silver,  determination  of 

amount    163 

in  base-metal  solutions  162, 
246 

in  Stetefeldt  furnace . .     85 
in  sulphating  roasting    97 
Lucky  Tiger  mine,  Sonora,  Mex- 
ico, experiments  with  ore  from    36 

Lump-grinding  machine 238 

Lye,     manufacture     of,      from 
wood  ashes 205 

McDougal  furnace,  fuel  needed  in    42 
Mansfeld,  Germany,   process  of 
sulphating  roast- 
ing at     95 

Ziervogel  process  at    94, 
281 

Mechanical  roasting  furnace  ...  62 
fed  by  charges ...  63 
vs.  reverberatory 

hand  worked  . .     62 
with      continuous 

feeding 71 

Hofmann's    im- 
proved How- 
ell  furnace  . .     81 
Howell-White 

furnace 77 

O  'Harra  furnace     71 
Ropp  furnace. .     74 
Stetefeldt     fur- 
nace       82 

Mercury,  action    of    base-metal  ' 

chlorides  on 40 

effect  of  amalgamation 

on  37 

Metal  chlorides,     formation     of 
volatile,  the  cause  of 

loss  of  silver    30 

formed  in  roasting  ....  3,  4 
reaction  for  formation  of  4 
silver  chloride  soluble  in  40 
volatile,  action  of,  in 

chloridizing  silver  ...       4 
see  also  Base-metal  chlo- 
rides 


338 


INDEX 


Metal  oxides,  formed  in  roast- 
ing         8 

silicates,  action  of  steam 

on,  in  roasting    31 

subchlorides     formed     in 

heap-roasting    30 

sulphates,       formed       in 

roasting 3 

sulphides,  effect  of,  on  free 

percolation    12 

Mexican  Santa  Barbara  Mining 

Company    61 

Milk  of  lime,  as    precipitant    of 

silver    164 

used    in    preparing 
calcium    polysul- 

phide    186 

Modified  Ho  well  furnace,   com- 
pared with  reverber- 

atory    126 

cost  of  roasting  in. .    119 
results     of     experi- 
ments with. .  .116, 120 
roasting  San   Fran- 
cisco del  Oro  ore 

in 113,243 

Mohr's  burette 327 

Monitor,  California,  lixiviation  at  155 

Muffle,  experiments  in   143, 196 

loss  of  silver  in  samples 
roasted  in    20 

Native  silver,  in  ore  from  San 

Francisco  del  Oro  mine   100 

Nitric  acid,  used  in  test  for  silver  183 
North    Mexican    Silver    Mining 
Company,  Mexico 69, 240 

O'Harra  furnace   51,  71 

Oker,    Germany,    extraction    of 
silver  from  copper  matte  from  278 

Oker  process    278 

Ontario,  Utah,  roasting  ores  from    41 
Ores,  silver,  classification  of  in 

relation  to  roasting    ...       9 
suitable    for    chloridizing 
roasting 3 


Oxide  of  antimony,  formed  in 

roasting 7 

Oxidizing  period  of  roasting  /. .       4 
roasting,  experiments 
with  two-story  re- 
verberatory    fur- 
nace      122 

necessary  for  cer- 
tain ores 16 

Oxnam,  T.  H 288 

Oxygenation  of  gold  ores 296 

Palmarejo  and     Mexican     Gold 
Fields,    Ltd.,    Chi- 

nipas 288 

Chihuahua,  Mexico    .   288 

ores 296 

Pan  amalgamation 37 

evaporator 270 

Parral,  Chihuahua,  Mexico,  11,  170, 

174 

Patio 291 

Pearce  turret  furnace  ...  71,  261, 263 

Pelton  wheels 289,  304 

Percy    155 

Physical  changes  in  roasting  ore     13 

Plattner    85 

Plattner's  method 283,  285,  324 

Plomosas,  Mexico 41 

Plumbiferous  silver  ores,  lixivia- 

ation  of   220 

roasting   41 

Porphyry,  behavior  of,  in  roasting       8 

Potassium  cyanide 287 

solution  used  to  ex- 
tract silver 104 

Precipitant,  adding,  in  precipi- 
tation of  silver 191 

Precipitate,   pressure   tanks  for 

treatment  of.  ...   205 
refining  by  cupella- 

tion 195 

removal  of  sulphur 

from 203 

silver,  drying  and 
roasting  furnace 
for  . .  .  208 


INDEX 


339 


Precipitate,  silver,  refining  the     208 
treatment  of..  .194,198 

Precipitation   320 

of  silver 185 

adding  the  pre- 
cipitant     191 

and  gold 299 

chloride    by  di- 
lution      with 

.water 170 

from  base-metal 

solutions ....    164 
preparing      cal- 
cium     hypo- 
sulphide  for  .    186 
process,     descrip- 
tion of 156 

tanks    156 

vats    185, 233 

Pressure  tanks 199,  200,  205,  265 

Producer  gas,  as  fuel  for  roasting     42 
Purifying  tower .   267 

Quartz,  a    desirable    gangue    in 

roasting     8 

behavior  of,  in  roasting       8 
in  gangue  of  ore  from 
San  Francisco  del  Oro 
mine 100 

Refining  the  silver  precipitate.  .  208 

Reverberatory  furnace  ....  143,  254, 

263, 278 

compared  with  Bruckner    30 

description  of 45 

dust  formed  in   87 

experiments     with      in 

heap-roasting 29 

fuel  needed  in,  for  roast- 
ing ores  poor  in  sul- 
phur        43 

fumes    not    visible    in 

roasting  in 30 

hand-worked,    vs.    me- 
chanical roasting.  ...     62 
long,  see  Long  reverber- 
atory  furnace 


Reverberatory    furnace,   loss  of 
silver  in,  greater  than 
in  Bruckner  furnace  .     20 
roasting  calcareous  ore 

in    144 

roasting      plumbiferous 

silver  ores  in 41 

rules    for    temperature 

and  draft  in 149 

single-hearth 45 

two-story  long 60 

two-story  single-hearth  46 
used  at  Sombrerete. ...  31 
used  to  remove  sulphur 

from  precipitate  ....  207 
Rising  Star  mine,  Flint,  Idaho, 

ore  from 72 

Roasting,  appearance  of  ore  in 

steps  of    5 

auriferous  silver  ore  .   323 
chloridizing,  see  Chlo- 
ridizing  roasting 

heap-    27,31 

self- 26,65 

for  amalgamation  ...     37 
in  the  Bruckner  fur- 
nace      128 

methods  of    26 

silver,  ores  containing 

gold 35,  285 

steps  in  process  of   . .       4 

sulphating 94 

Rolls 11,143 

Roof  of  long  reverberatory  fur- 
nace, hight  of 53 

Ropp,  Alfred  von  der 75 

Ropp  furnace 51,  74 

Russell    163, 166, 178,  253 

Russell  method  of  precipitating 

lead 179 

Russell  process    251 

Russell's  extra  solution,  used  to 
extract  silver 104,  252, 254 

Salt,  action  of,  in  roasting 4 

adding  during  crushing  pro- 

19 


340 


INDEX 


Salt,  adding  in  battery 140 

addition  of,  in  amalgama- 
tion       38 

best  time  to  add,  in  roasting     16 
coarse    vs.    pulverized    in 

roasting 18 

effect  of  adding  in  roasting      9 
excess  of,  on  loss 

of  silver 164 

in  roasting  ore 
from  San  Fran- 
cisco del  Oro 

mine 112 

in  roasting  zinc- 
lead  ore  102 

on  balling  of  ore.   142 
on  calcareous  ores  135 
percentage    of,    needed    in 

roasting 15 

pulverized    vs.    coarse    in 

roasting 18 

San  Francisco  del  Oro  ore    17 

analysis  of  raw 100 

analysis  of  roasted 118 

chlorination  after  ore  has 

left  furnace 115 

conclusions     of     experi- 
ments in  chloridizing  111 
cupric   chloride   used   in 

leaching 174 

effect  of  salt  on 103 

experiments     in     chlori- 
dizing   31,99 

crushing 11 

heap-roasting 28 

trough  lixiviation  . . .  243 
furnaces  used  in  roasting  61 
loss  of  silver  in  leaching  162 

in  roasting 117 

methods  of  treating  ....  252 
results     of    roasting    in 
modified  Howell  fur- 
nace      115 

salt  required  in  roasting.  16 
stock  solution  used  on  . .  175 
time  required  to  work  by 

lixiviation   ...          .  248 


San  Francisco  del  Oro  ore,  using 

Howell  furnace  on  . .     81 
wood  consumed  in  roast- 
ing     119 

San  Marcial,  Mexico 155 

San  Salvador,  Central  America.   321 
Sand,  as  filter  in  silver  leaching  182 

cyanidation  of 292 

-retaining  tank 290 

Santa  Barbara 11 

Schemnitz,  Hungary,  Ziervogel's 

method  tried  at 95 

Self-roasting  calcareous  ores   ...    140 

chloridizing 26,  65 

Settling-tank  for  sluicing 225 

Silver  chloride  decomposed     by 

caustic  lime 8 

obtained  by  chloridizing 

roasting.  / 3 

precipitating  with  water  170 
soluble    in    solution    of 

metal  chlorides 40 

solubility  of 161 

copper  glance,  behavior  of, 

in  roasting 9 

extraction  of,  by  the  Zier- 

vogel  process 281 

from  black  copper  ....   278 
from  copper  matte  ....   258 
with  sulphuric  acid   . . .   258 
Hofmann's  method  of  ex- 
traction with  sulphuric 

acid 259 

leaching,  calcium  hyposul- 
phite, action  of  179 
calcium  sulphide 
as    precipitant 

in    179 

description  of   . .    156 
effect  of  lead  and 

copper   178 

end  of 183 

filters  for  leach- 
ing tanks  ....    182 
in  lixiviation  with 
soduim    hypo- 
sulphite        174 


INDEX 


341 


Silver    leaching  in  ores  rich  in 

gold 284 

in  trough  lixivia- 

tion 246 

method  of   177 

regeneration      of 
hypo  solution     181 

tanks  for   229 

testing  for  silver 

in    178 

time  required  180,248 
loss  of,  by  volatilization  20, 117 
determination     of 

amount    163 

in  base-metal  solu- 
tion    246 

in  roasting,  meth- 
ods of  ascertain- 
ing    22 

in  Stetefeldt  fur- 
nace    85 

in  sulphating  roast- 
ing    97 

reduced  by  steam  33 
methods  of  chlorination  of  3 
native,  in  ore  from  San 

Francisco  del  Oro  mine  100 
ores,  classification    of    in 

relation  to  roasting      9 
rich  in  gold,  cyani- 

dation  of   287 

treatment  of 283 

precipitate,  fineness  of  . .  248 
treatment  of . .   194 

precipitation  of 185 

from    base-metal    solu- 
tions     164 

in  cyaniding    299 

prevention  of  loss  of    ....    162 
recovery    of    from    waste 

liquor 165 

testing    cyanide    solution 

for 326 

Silver  King  mine,  Arizona,  Bruck- 
ner furnace  mod- 
ified to  roast 
ores  from 67 


Silver  King  mine,  desilverizing 
base-metal  chlo- 
rides at 170 

experiments    with 

ore  from  ....'..  33 
leaching  with  cu- 

pric  chloride  at  174 
lixiviation  at  ....  155 
regenerating  hypo 

solution  at  ....  181 
time  required  for 

lixiviation  at  . .   180 
Single-hearth  reverberatory  fur- 
nace        45 

two-story 46 

Sizing-test  on  slime 305 

Slate,  behavior  of,  in  roasting. .       8 

Slime,  formed  in  crushing 12 

method  of  treatment  ...   313 

pits 291 

plant,  description  of    ...   305 
settling  rate  per  hour    . .   314 

sizing  test  on 305 

treatment  of,  in  cyanid- 
ing   auriferous    silver 

ore    304 

Sluice-tanks  225 

Sluicing 225 

Sodium  carbonate,  precipitate 

for  lead    178,179 

chloride,  addition  of,  to 

ore    3 

formed  in  roasting  .  .       4 
method  of  decompo- 
sition of,  in  roast- 
ing         3 

cyanide    287 

hyposulphite,  a  solution 

for  arsenate  of  silver      7 
action     of,    in    chlo- 
ridizing      zinc-lead 

ores 104 

best  strength  of  solu- 
tion    177 

handling    201 

in  silver  leaching  ...   174 
lixiviation  with  ...    .   155 


342 


INDEX 


Sodium  hyposulphite,  regenera- 
tion of,  when  foul.  181 
used  in  extracting  gold 

from  silver  ore ....     35 
used  in  leaching  zinc- 
lead  ore 106 

used    on     calcareous 

ores 127 

silicate,  formed  in  roast- 
ing    8 

sulphate,  accumulation 
of,  in  leaching  solu- 
tion    197 

cause  of  balling 143 

formed  in  roasting  . .       4 

sulphide,  as  precipitant    164, 

171 

as  test  for  silver 178 

Solubility  of  silver  chloride  ....   161 

Solution  tanks     312 

Sombrerete,  Zacatecas,  Mexico    254 
base-metal  leaching  at  ...    166 

experiments  at 30 

experiments  in  heap-roast- 
ing ore  from  28 

loss  of  silver  in  leaching 

ore  from 163 

ore  from  11, 17 

reverberatory     furnaces 

used  at 31 

straw  for  filter  used  at  . .    182 
two-story  long  furnaces  at     61 
Stamp  battery,  adding  salt  in.   140 
effect      of,      in 

crushing  ore       11 

Starch  paper,  test  for  silver. .  .    178 
Steam,  applied  in  Howell  furnace  110 
purifying  tower .  267 
roasting  by  Von 

Patera 155 

chloridizing  roasting  with     31 
consumption   of    fuel   in 

use  of,  in  roasting  ...  32 
effect  of,  in  roasting  ....  32 
in  base-metal  leaching  at 

Sombrerete   169 

loss  of  silver  reduced  by     33 


Steam  pump,    used    instead    of 

pressure  tanks 199 

used  in  desilverizing  waste 

liquor 165 

used  in  roasting  to  aid 

the  extraction  of  silver      1 3 
used  in  extraction  with 

sulphuric  acid 259 

Stetefeldt,  C.  A.,  27,  41,  83,  128, 166, 

208 

Stetefeldt  furnace,  capacity  of.     85 
conclusions   of   ex- 
periments on  zinc- 
lead  ores  in  ....    Ill 
description  of  ....     82 
draft  required  in..    109 
dust  formed  in.  ...     87 
experiments  in  re- 
roasting      ore 
from  shaft  of  ...    107 
experiments  with  .     31 
in  heap-roasting  28,  29 
failure   on    ores  of 

Mexico 86 

fuel  needed  in  ....     43 
in  Bosque   mill   at 

Parral 101,102 

preferable    to   long 

reverberatory ...     52 
silicates  formed  in.    110 
unsuitable  for  zinc 
blende    and    ga- 
lena ores    Ill 

used     at     Sombre- 
rete,    Zacatecas, 

Mexico 254 

used     in     roasting 
*  plumbiferous  sil- 
ver ores    41 

used  to  roast  zinc- 
lead  ore 104 

Stir  tanks 263 

Storage  tanks 298 

for  hyposulphite 

solution    203 

Straw,  used  for  filter  in  leaching 
tanks  . ,  .182 


INDEX 


343 


Sulphate    of    lime,    formed    in 

roasting 8 

Sulp hating  roasting 94 

time  required  for.     97 
Sulphide  minerals,  classification 
of,     in    relation    to 

roasting 9 

ores 3 

Sulphur  in    ores,    effect    of,    on 

quantity  of  fuel    ....     42 
lack    of,    overcome    by 

burning  brimstone  .  .     86 
removal  of,  by  burning    207 

by  distillation    207 

from      silver      precipi- 
tate     203 

used   in   preparing   cal- 
cium polysulphide   .  .    186 
Sulphuric  acid,  anhydrous,     for- 
mation of  ....       4 
extraction  with  .   258 
formed     in     sul- 
phating   roast- 
ing         94 

gas,     formed     in 

roasting 4 

Sulphurous  acid  converted    into 
anhydrous    sul- 
phuric acid       4 
formed  from  ga- 
lena in 

roasting  .  .        6 
in  roasting  .  .        3 
gas,    formed    in 
heap-roasting    28, 
29 
chloride,    formed    in 

roasting 3,6,7 

Sump-tanks 298 

Sustersic,  F 8, 195, 197,  204 

Sustersic's  method  of  preparing 
precipitate  for  refining 195 

Tailing  elevator-wheel   290 

Tank,   leaching,  construction  of  157 

lixiviation    219 

fineness  of  precipitate .   248 


Tank  lixiviation   less   advanta- 
geous than  trough  lixi- 
viation      249 

quantity  of  solution  re- 
quired     248 

time  required  for  base- 
metal  leaching.  245 
for  silver  leaching  .  .   248 
Tarshish  mine,  Alpine  county, 

California 283,  285 

Testing  cyanide  solution  for  gold 

and  silver 326 

Tests  for  silver,  in  leaching. ...    178 
Tools,  best  form  of,  for  working 

charges 52 

Tower  for  refining  cupric   sul- 
phate solutions 267 

Trinidad,     Mexico,     lixiviation 

at 155 

Trough,  distributing,  for  milk  of 

lime    188 

lixiviation   181,  219 

advantages  of 249 

arrangement    and 

operations  of  tanks  229 
at  Cusihuiriachic  ....  240 
fineness    of     precipi- 
tate     248 

precipitating  vats.  . . .   233 
prevention  of  loss  of 

silver  in 162 

quantity   of    solution 

required 248 

settling-tank 225 

silver     dissolved     by 
base-metal  solution 

in    246 

silver  leaching    246 

sluice-tanks  and  sluic- 
ing       225 

time  required  for  base- 
metal  leaching  ....   244 
time  required  for  silver 

leaching 248 

water  required   245 

Troughs,  lixiviation    215 

Two-story  long  furnace    60 


344 


INDEX 


Two-story  reverberatory  furnace 
compared  with  mod- 
ified Howell 126 

consumption  of 

wood  in   125 

cost  of  roasting  in.    126 
experiments    with, 
in    roasting    San 
Francisco  del  Oro 

ore 121 

used  in  sulphating 

roasting 95 

single-hearth     rever- 
beratory furnace  .  .     46 

Underfed  stoker    272 

United  Zinc  and  Chemical  Com- 
pany, Argentine,  Kansas  ....     90 

Veta  Grande,     Parral,     Mexico, 
leaching  with  cupric 

chloride  at 174 

mine 101 

Volatile    chlorides,    method    of 

avoiding  expulsion  of 20 

Volatilization 20 

loss  of  silver  by    .   117 
Von  Patera 155, 179 

Waste  liquor,  recovery  of  silver 

from    165 

Water,  consumption  of,  in  trough 

lixiviation    245 

use  of,  to  desilverize  base- 
metal  chlorides 170 

White-Howell  furnace,  at  Parral, 

Mexico 170 

roasting      zinc- 
lead  ore  in . .    1 12 
see  also  Howell- 
White  furnace 
White  lead,  not  to  be  used  in 

lixiviation  troughs    215 

Wilfley  concentrators 290 

Wood,  amount  required  in  roast- 
ing zinc-lead  ore 109 

ashes,  making  lye  from.  205 


Wood,  consumption  of,  in  rever- 
beratory furnace  125, 150 
in  roasting  San  Fran- 
cisco del  Oro  ore  ...  119 

in  self-roasting 141 

used  as  fuel  in  roasting.  42 

Woolly  ore,  cause  of 13 

Working  doors,  construction  of, 

in  long  reverberatory  furnace  56 

Yedras,  Sinaloa,  Mexico,  behav- 
ior of  ore  from  ...  17 

conclusions  of  experi- 
ments with  ores  at  151 

loss  of  silver  in  roast- 
ing ore  from 21 

ores   from 50 

roasting  calcareous 
ores  at 145 

tests  in  loss  of  weight 
in  roasting  ores 
from 23 

Ziervogel 94 

process 94,281 

Zinc  blende,  behavior      of       in 

roasting 6,9 

effect  of  salt  on,  in 

roasting 16 

steam  on  ores  con- 
taining       33 

in  ore  from  San 
Francisco  del  Oro 

mine 99 

not  favorable  in 
roasting  in  Stete- 
feldt  furnace  ...  86 

boxes   299,312 

clean  up  of    300 

chloride,  formed  in  roasting      7 
fumes,  effect  of,  in 

roasting 7 

effect  of,  in  ores 15 

effect  of,  on  time  of  lixivia- 
tion      180 

-lead  ore, argentiferous,  chlo- 

ridizing  of   ....     99 


INDEX 


345 


PAGE 

Zinc-lead    ore,    chloridizing    in 

Howell  furnace      82 
from    San     Fran- 
cisco    del     Oro 
mine 28,31,61 


PAGE 

Zinc  oxide,  formed  in  roasting .         7 
sulphate,  formed  in  roast- 
ing           7 

sulphide,  action  of,  in  roast- 
ing           6 


YD  07559 


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